Influence of temperature and time on reduction behavior in iron ore–coal composite pellets

Dealkalization processes of bauxite residue: A comprehensive review

The critical applications of vanadium in metallurgical field and the growth in commercialization of vanadium redox flow batteries (VRFB) have led to the increased demand of vanadium. It is thus important to ensure the sustainability of vanadium production. Vanadium bearing slags, the solid byproducts in iron- and steel-making plants, are the principal source of vanadium production, accounting for more than 69% of total vanadium in terms ofthe raw material types. Academic researches and engineering investigations have been addressed to develop such metallurgical processes for treating the vanadium bearing slags. This article presents a comprehensive review on the metallurgical treatments of vanadium bearing slags. The composition and phase/mineralogical characterization of vanadium bearing slags from various sources are given. Literature review shows that the vanadium bearing slags have been traditionally treated through the roasting-assisted leaching with the recent efforts of integrating the state-of-the arts technologies in extractive metallurgy and developing direct leaching methodologies. Some promising methods are worth discussing and quite encouraging, and expected to be the future focuses of this area. Discussion also highlights the separation of vanadium from silica, phosphorus and chromium as the major interfering elements/metal in the leach solutions of vanadium slags. Recommendation is made for taking up future works in order to develop a sustainable metallurgical process for vanadium bearing slag.

A review on the metallurgical recycling of vanadium from slags: towards a sustainable vanadium production

A semi-industrial experiment of suspension magnetization roasting technology for separation of iron minerals from red mud.

Evaluation of red mud as a polymetallic source – A review

A process with coal-based direct reduction followed by magnetic separation is presented for recovering metallic iron from high-phosphorus oolitic hematite in this study. Ca(OH)2 and Na2CO3 were used as additives in the reduction roasting. A Direct Reduction Iron (DRI) with 93.28 mass% Fe, and 0.07 mass% P can be obtained at a recovery percentage of 92.30 mass% under optimal conditions. The mechanisms of Ca(OH)2 and Na2CO3 were investigated by XRD and SEM with EDS. It showed that fluorapatite was reduced to P smelt into metallic iron without additives while the hematite was reduced. The addition of Ca(OH)2 can not only inhibit the reduction of fluorapatite but also promote the reduction of hematite. Na2CO3 can promote the separation of iron from slag, meanwhile it may also inhibit the reduction of fluorapatite at the presence of 15 mass% Ca(OH)2. Under optimal conditions, phosphorus remained as fluorapatite in the slag and can be removed by grinding and magnetic separation.

The Function of Ca(OH)2 and Na2CO3 as Additive on the Reduction of High-Phosphorus Oolitic Hematite-coal Mixed Pellets

The effect of coal levels on phosphorus removal from a high phosphorus oolitic hematite ore after direct reduction roasting have been investigated. Raw ore, coal, and a dephosphorization agent were mixed and the mixture was then roasted in a tunnel kiln. The roasted products were treated by two stages of grinding followed by magnetic separation. XRD and SEM–EDS examination of the products was used to analyze differences in the roasted products. The results show that coal is one of the most important factors affecting the direct reduction roasting process. When the inner coal levels increased from 0% to 15% the iron grade decreased linearly from 94.94% to 88.81% and the iron recovery increased from 55.94% to 92.94%. At the same time the phosphorus content increased from 0.045% to 0.231%. Increasing the inner coal levels also caused more hematite to be reduced to metallic iron but the oolitic structure of the roasted product was preserved in the presence of high coal loading. The phase of the phosphorus in raw ore was not changed after direct reduction roasting. The effect of coal on the phosphorus content in the H-concentrate arises from changes in the difficulty of mechanically liberating the metallic iron from the phosphorus bearing minerals.

Effect of coal levels during direct reduction roasting of high phosphorus oolitic hematite ore in a tunnel kiln

A novel utilization of Bayer red mud through co-reduction with a limonitic laterite ore to prepare ferronickel

In this article, mineralogical phase changes and structural changes of iron oxides and phosphorus-bearing minerals during the direct reduction roasting process were investigated by X-ray diffraction (XRD) and scanning electron microscope (SEM) It has been found that the reduction of hematite follows the following general pathway: Fe2O3 → Fe3O4 → FeO → Fe The last step of the reduction process contains two side reactions: either FeO → Fe2SiO4 → Fe or FeO → FeAl2O4 → Fe depending on the micro mineralogical makeup of the ore In the reduction process of FeO → Fe, oolitic structure was destroyed completely and fluorapatite was diffused into gangue while metallic phase is coarsening at temperatures below 1200°C Therefore, the separation of phosphorus-bearing gangue and metallic iron can be achieved by wet grinding and magnetic separation, and low phosphorus content metallic iron powder can be obtained However, when the temperature reached 1250°C and beyond, some of the fluorapatite was reduced to elemental

Study on Phosphorus Removal of High-Phosphorus Oolitic Hematite by Coal-Based Direct Reduction and Magnetic Separation

Embedding direct reduction followed by magnetic separation was conducted to fully recover iron and titanium separately from beach titanomagnetite (TTM). The influences of reduction conditions, such as molar ratio of C to Fe, reduction time, and reduction temperature, were studied. The results showed that the TTM concentrate was reduced to iron and iron-titanium oxides, depending on the reduction time, and the reduction sequence at 1200 °C was suggested as follows; Fe 2. 75 Ti 0. 25 O 4 →Fe 2 TiO 4 →FeTiO 3 →FeTi 2 O 5 . The reduction temperature played a considerable role in the reduction of TTM concentrates. Increasing temperature from 1100 to 1200 °C was beneficial to recovering titanium and iron, whereas the results deteriorated as temperature increased further. The results of X-ray diffraction and scanning electron microscopy analyses showed that low temperature (≤ 100 °C) was unfavorable for the gasification of reductant, resulting in insufficient reducing atmosphere in the reduction process. The molten phase was formed at high temperatures of 1250–1300 °C, which accelerated the migration rate of metallic particles and suppressed the diffusion of reduction gas, resulting in poor reduction. The optimum conditions for reducing TTM concentrate are as follows: molar ratio of C to Fe of 1. 68, reduction time of 150 min, and reduction temperature of 1200 C. Under these conditions, direct reduction iron powder, assaying 90. 28 mass% TFe and 1. 73 mass% TiO 2 with iron recovery of 90. 85%, and titanium concentrate, assaying 46. 24 mass% TiO 2 with TiO 2 recovery of 91. 15%, were obtained.

Chengyan Xu

Papers

The Function of Ca(OH)2 and Na2CO3 as Additive on the Reduction of High-Phosphorus Oolitic Hematite-coal Mixed Pellets

Effect of coal levels during direct reduction roasting of high phosphorus oolitic hematite ore in a tunnel kiln

A novel utilization of Bayer red mud through co-reduction with a limonitic laterite ore to prepare ferronickel

Study on Phosphorus Removal of High-Phosphorus Oolitic Hematite by Coal-Based Direct Reduction and Magnetic Separation

Effects of embedding direct reduction followed by magnetic separation on recovering titanium and iron of beach titanomagnetite concentrate