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.

Simple magnetic separation for a certain magnetite mine with ultrafine disseminated lean ores has resulted in low performance, as the fine sizes and aggregation of ground mineral particles have caused inefficient recovery of the ultrafine minerals. In this study, we attempt to increase the apparent sizes of target mineral particles, and improve the separation indices, by using a multi-stage grinding-dispersion-selective flocculation-weak magnetic separation process. The results showed that under the conditions of 500 g/t sodium hexametaphospate (SHMP) as dispersant, 750 g/t carboxymethyl starch (CMS) as flocculant, agitating at 400 rpm for 10 min, with slurry pH 11, and final grinding fineness of 93.5% less than 0.03 mm, the obtained concentrate contained 62.82% iron, with recovery of 79.12% after multi-stage magnetic separation. Compared to simple magnetic separation, the concentrate’s iron grade increased by 1.26%, and a recovery rate by 5.08%. Fundamental analysis indicated that, in a dispersed state of dispersion, magnetite particles had weaker negative surface charges than quartz, allowing the adsorption of negative CMS ions via hydrogen bonding. Consequently, the aggregate size of the initial concentrate increased from 24.30 to 38.37 μm, accomplishing the goal of selective flocculation, and increasing the indices of separation.

How to utilize low grade complex iron resources is an issue that has attracted much attention due to the continuous and huge consumption of iron ores in China. High-aluminum iron ore is a refractory resource and is difficult to upgrade by separating iron and alumina. An innovative technology involving synergistic reducing and synergistic smelting a high-aluminum iron ore containing 41.92% Fe_total, 13.74% Al₂O_3, and 13.96% SiO₂ with a high-manganese iron ore assaying 9.24% Mn_total is proposed. The synergistic reduction process is presented and its enhancing mechanism is discussed. The results show that the generation of hercynite (FeAl₂O₄) and fayalite (Fe₂SiO₄) leads to a low metallization degree of 66.49% of the high-aluminum iron ore. Over 90% of the metallization degree is obtained by synergistic reducing with 60% of the high-manganese iron ore. The mechanism of synergistic reduction can be described as follows: MnO from the high-manganese ore chemically combines with Fe₂SiO₄ and FeAl₂O₄ to generate Mn₂SiO₄, MnAl₂O₄ and FeO, resulting in higher activity of FeO, which can be reduced to Fe in a CO atmosphere. The main products of the synergistic reduction process consist of Fe, Mn₂SiO_4, and MnAl₂O₄.

The magnetic differences of the artificial magnetite and the natural magnetite were studied to improve the magnetic parameters of artificial magnetite in the study. The effect of roasting conditions on the magnetism of artificial magnetite and the diffidence of magnetism and magnetic behavior of the artificial magnetite and natural magnetite was studied and compared. The results indicate that the magnetically strong artificial magnetite has been obtained under the optimized process parameters of the roasting temperature of 800°C, roasting time of 60min and the mass ratio of pulverized coal to ore (C/O ratio) of 2%. And the magnetic analysis shows that the magnetism of natural magnetite is stronger than that of artificial magnetite under the same conditions, but the remanence and coercivity of artificial magnetite is stronger than natural magnetite. With the decrease of particles, the magnetism of the artificial magnetite changes insignificantly, but the magnetization of natural magnetite decreases gradually; and the difference in the magnetism of two becomes smaller (from 1.53 × 10^-4·m³·kg^-1 to 0.14 × 10^-4m·³·kg^-1) with the decrease of particles. The gaps and impurities in the ore are the cause of the difference, which provides a theoretical basis for the further study the size-fractionated magnetic separation of artificial magnetite.

Magnetization reduction roasting is an important method for the utilization of oolitic magnetite. In this study, the magnetization reduction behavior and kinetics of oolitic hematite in gas-based roasting were systematically investigated by X-ray photoelectron spectroscopy (XPS). The results revealed that under optimal roasting conditions of 650 °C, a roasting time of 60 min, and a CO concentration of 30%, the magnetization reduction rate of the roasted product reached 44.34%. Furthermore, the weak magnetic separation concentrate presented a TFe of 58.09% and a concentrate iron recovery of 94.3%. The results of the XPS spectrum indicated that the peak area ratio (Fe²⁺/Fe³⁺) gradually increased with an increase in roasting temperature, roasting time, and CO concentration, while over-reduction occurred when the roasting temperature exceeded 750 °C. The investigation of magnetization roasting kinetics for varying particle sizes demonstrated that the magnetization reduction process is controlled by chemical reaction, with a corresponding activation energy range of 42.96 kJ/mol to 63.29 kJ/mol, indicating the particle size has little effect on the magnetization reduction of oolitic hematite.

Sintering, as one of the major sources of carbon emissions in the steel industry, is in urgent need of innovation in carbon reduction technology under the targets of carbon peaking and carbon neutrality in China. We propose the carbon free sintering process which is similar to the belt roaster, where heat is provided by gas fuel combustion for the layer. In order to achieve the process, this study uses the micro-sintering method to simulate it by varying the sintering temperature, sintering time and binary basicity of the mixture. It was investigated the process of mineralisation solidification of carbon-free sintering and to provide relevant guidance for future carbon-free sintering process experiments. As the sintering temperature increased, the strength and calcium ferrite of the samples first increased and then decreased, reach the maximum value at 1300°C and 1275°C respectively. However, the porosity of the samples first decreased and then increased, reach the minimum value at 1275°C. As the sintering time increased, the strength and calcium ferrite first increased and then decreased, reach the maximum values at 12 min and 6 min respectively, and the porosity of the samples first decreased and then increased, reach the minimum values at 6 min. As the binary basicity increased, the strength of the samples first decreased and then increased, reach the minimum value at basicity 1.8, and the calcium ferrite and porosity of the samples gradually increased. The carbon-free sintering process does not change the consolidation process of iron ore, and it has theoretical feasibility.