Pyroelectrometallurgical processing of bismuth-containing oxides

In this work, we substantiate and develop a general pyroelectrometallurgical technology for processing bismuth dross and oxides (the intermediate products of lead bullion refining by the Betterton-Kroll process) to obtain crude bismuth. The research focuses on bismuth dross (3–5% Bi; 80–85% Pb) remelted at 500–600°С in the presence of NaNO3 and NaOH, as well as the obtained alkaline melt (bismuth oxides, 1–5% Bi; 60–70% Pb). The conducted experiments allowed us to determine optimal parameters of the main steps of processing bismuth oxide, as well as the characteristics of obtained products. Reduction smelting of bismuth oxides at 1150°C (with the addition of sodium carbonate, quartz and fine coke in the amount of 66, 25 and 5% of bismuth oxides mass, respectively) is proposed, leading to bismuth lead formation. Its decoppering is carried out at 350–600°C with 2.0% sulfur (by its weight), added to the melt. We propose to carry out the alkaline treatment of the decoppered Pb-Bi alloy at 500oC in contact with sodium hydroxide, sodium nitrate and sodium chloride, taken in amounts up to 10.2, 8.3 and 1.4% by weight of bismuth lead, respectively. Subsequent electrolysis comprises electrolytic processing of purified Pb-Bi alloy ingots at 550oC. The electrolyte consists of a melt with the following composition, %: NaCl – 7, KCl – 35, PbCl2 – 18 and ZnCl2 – 40. As a result, two end products were obtained by the proposed bismuth oxide processing. The anodic product at the second stage of electrolysis, crude bismuth (yielded 1.1% by the weight of oxides) contains 93.62% Bi and 4.14% Pb, extraction from oxides amounts to 19.0% Bi and 0.1% Pb. About 1.2% Bi and 9.1% Pb of their initial content in the oxides are transferred to the cathodic product containing 0.033% Bi and 97.83% Pb (the yield equalled 5.1% of the oxides).

Materials and Energy Balance of E-Waste Smelting—An Industrial Case Study in China

The application of Nerin Recycling Technologies (NRT) in electronic waste (E-waste) smelting was introduced in this study, and the material and energy balance was calculated based on the practical data with the METSIM software (METSIM International, USA). The main results are as follows: (1) the optimized processing parameters in the NRT smelting practice were the E-waste feeding rate of 5.95 t/h, oxidation smelting duration of 3.5 h, reduction smelting duration of 0.5 h, oxygen enrichment of 21–40 vol.%, oxygen consumption of 68.06 Nm3/ton raw material, slag temperature of 1280 °C, slag composition: Fe/SiO2 mass ratio of 0.8–1.4, CaO, 15–20 wt.%, Cu in crude copper ≥ 95 wt.%, Cu in slag, 0.5 wt.%, recovery of Cu, Au, and Ag ≥ 98%; (2) 98.49% Au, 98.04% Ag, 94.11% Ni, and 79.13% Sn entered the crude copper phase in the smelting process, 76.73% Pb and 67.22% Zn volatilized to the dust phase, and all halogen elements terminated in the dust and off-gas; (3) total heat input of the process was 79,480 MJ/h, the energy released by chemical reactions accounted for 69.94% of the total, and heat from fuels burning accounted for 33.04%. The energy brought away by the off-gas was 38,440 MJ/h, which was the largest part in heat output. The heat loss with the smelting slag accounted for 28.47% of the total.

The Recovery of Metallic Tin from an Industrial Tin-Bearing By-Product Containing Na2SO4 by Reduction Smelting Process

Tin was recovered in metal from an industrial tin-bearing byproduct containing Na2SO4 by carbothermic reduction smelting, and the effects of basicity (Na2O/SiO2), temperature, and reaction time on the recovery of tin were studied. Na2SO4 was reduced by carbon and formed into sodium silicate slag (Na2O–SiO2) in the presence of SiO2. Tin content in slag decreased with the increase of Na2O/SiO2 ratio in slag, temperature, and reaction time, but the recovery of tin was affected by volatilization of tin in high temperature and high silica region of basicity. In this study, the maximum recovery rate of tin was 94.8% at the experimental condition of 1200 °C, 2 h, and 0.55 of Na2O/SiO2 ratio. The major impurities in produced metal were Bi, Pb, Cu, Fe, and most of Bi, Pb, Cu were distributed to the metal phase, but the distribution of Fe was closely related to basicity.

Identification of patterns in the structural and phase composition of the doping alloy derived from metallurgical waste processing

This paper reports a study into the structural-phase composition of the doping alloy made by processing metallurgical anthropogenic waste involving reduction smelting. This is required for determining the technological parameters that ensure an increase in the level of extraction of target elements during the processing of anthropogenic waste and for the further use of the doping alloy. It was revealed that the phase composition of the doping alloy manifested a solid solution of the doping elements and carbon in α-Fe. Cementite Fe3C and silicides Fe5Si3, FeSi, and FeSi2 were also identified. In this case, the doping elements were more likely to act as substitution atoms. It has been determined that the microstructure of the alloy consisted of several phases of different shapes and contents of the basic doping elements. Sites with an elevated iron level of up to 95.87 % by weight in the composition could be represented by the solid solution phase of the doping elements and carbon in α-Fe. The sites with a relatively high (% by weight) content of carbon (0.83‒2.17) and doping elements ‒ W, up to 39.41; Mo, up to 26.17; V, to 31.42; Cr, to 9.15 ‒ were apparently of a carbide nature. The sites with a silicon content of 0.43‒0.76 % by weight likely included silicide compounds. The alloy's characteristics make it possible to smelt steel grades without strict carbon restrictions, replacing some of the standard ferroalloys. Neither phases nor compounds with a relatively high propensity for sublimation were identified in the material produced. Therefore, there is no need to provide conditions to prevent evaporation and loss in the gas phase of the doping elements. That could increase the degree of extraction of the doping elements

Sustainable and Clean Utilization of Yellow Phosphorus Slag (YPS): Activation and Preparation of Granular Rice Fertilizer

Yellow phosphorus slag (YPS) is a typical industrial solid waste, while it contains abundant silicon micronutrient required for the growth of rice. The key scientific problem to use the YPS as rice fertilizer is how to activate the slag efficiently during the phosphorite reduction smelting process. In this work, an alkaline rice fertilizer from the activated YPS was successfully prepared to use the micronutrients. Thermodynamic analyses of SiO2-CaO, SiO2-CaO-Al2O3, and SiO2-CaO-Al2O3-MgO systems were discussed to optimize the acidity for reduction smelting. Results showed that the reduction smelting followed by the water quenching process can realize the reduction of phosphorite and activation of YPS synchronously. Ternary acidity m(SiO2)/(m(CaO) + m(MgO)) of 0.92 is suitable for the reduction smelting and activation of the slag. After smelting, the molten YPS can be effectively activated by water quenching, and 78.28% P, 90.03% Ca, and 77.12% Si in the YPS are activated, which can be readily absorbed by the rice roots. Finally, high-strength granular rice fertilizers with a particle size of Φ2–4 mm were successfully prepared from the powdery nitrogen-phosphorus-potassium (NPK) and activated YPS mixture.

On the issue of complex processing of phosphorous iron ores

The reasons for the complete or partial rejection of the phosphorous iron ores (PIO) usage in ferrous metallurgy contrary to a shortage of iron ore raw materials are considered. Dephosphorization of metal from high-phosphorous cast irons leads to the complication of technology and deterioration of the production economics because of high intensity of the converter smelting and increasing requirements for the purity of steel. Limitation of PIO share in the blast furnace charge and decrease of phosphorus contents in cast irons led to a deterioration of the steel phosphate slag (PS) quality as fertilizers, which results in its disposal to the dump. Information on the opposite foreign practice of obtaining satisfactory phosphate slags is given, which is an increase in cast iron phosphorus contents by addition of ferrophosphorus. The reasons for abandoning this practice are analyzed. The authors studies aimed at increasing of phosphorus concentration in phosphate slags to bring PS conditions up to the requirements for specially produced by the chemical industry fused fertilizing phosphates and/or an intermediate product for the production of yellow phosphorus are described. On the basis of research and a technical and economic assessment, the feasibility of returning to the use of ferrophosphorus after making appropriate adjustments was substantiated. According to the proposed technology, phosphorous ferroalloy is used as a direct reagent in relation to substandard phosphate slags using their physical heat and exothermic reaction. The process is feasible outside the main metallurgical units and does not require the complication of steelmaking technology. It is possible to use ferrophosphorus as a by-product of the phosphorus industry due to its excess in the world market. It could also be obtained by reduction smelting of dump PS or current discharge slag. Phosphorous iron ores deep processing and the associated production of expensive chemical products will rise economic viability of “phosphorus-infected” iron ores usage.