China Nonferrous Metallurgy

Abstract:At present, Pidgeon process is used in magnesium production, which has high energy consumption, low resource utilization efficiency, serious environmental pollution and high carbon emission intensity. This paper analyzed the carbon emission status of magnesium smelting industry, and pointed out the ways of energy saving and emission reduction in magnesium smelting under the background of carbon neutralization from three aspects: calcining furnace, reduction furnace and new magnesium smelting technology. This paper promotes the development and application of new calcining furnace using heating ways of HTAC technology, oxygen-enriched combustion technology, etc., microwave, induction, plasma heating reduction furnaces and other new equipment, and develops new technologies with high efficiency, energy saving and carbon reduction, such as short-process green magnesium smelting technology, continuous smelting technology and alumino-thermic magnesium smelting technology, so as to finally achieve the goal of carbon neutrality and sustainable development of magnesium industry.

Abstract:The pyrometallurgical smelting process of copper sulfide ore is usually separated into concentrate smelting, copper matte converting and crude copper refining, among which smelting and converting can be further divided into suspension smelting and bath smelting according to process method. In recent years, bath smelting technologies have developed rapidly. This paper expounds on the development status, existing problems, solutions and schemes of oxygen-enriched bottom-blowing smelting, oxygen-enriched side-blowing smelting, bottom-blowing continuous converting and multi-lance top-blowing continuous converting in bath smelting. In particular, the prospects of various combination of continuous copper smelting processes are analyzed, continuous copper smelting technology and short process flow are the development trends of copper smelting in the future is pointed out.

Abstract:China owns cobalt reserves of around 80000t, accounting for about 1% of the global terrestrial cobalt reserves. However, it is also a major Co producer and consumer. In 2021, China’s refined Co output was 128000t, approximately 70% of the global total. Its Co metal consumption was over 80000t, more than 80% of which was for Li ion battery production. Therefore, China mainly relies on importing Co material, with a total of 90000t Co metal imported in 2021. The main raw materials of cobalt smelting are cobalt oxides, cobalt sulfides and cobalt alloy. In this paper, the characteristics and equipment for the smelting processes of the three kinds of materials were explained and summarized. On this basis, the main problems of China’s Co smelting were pointed out, namely excessive production capacity, overdependence on raw material imports, and to-be-improved product quality. The following suggestions were put forward: the development of Co industry shall be planned, orderly and scientific, and the blind expansion of Co smelting capacity shall be restricted; more R&D resources shall be committed to the recycling technologies of secondary Co resources to develop the circular economies of Co battery waste material, waste alloy, and waste catalyst; the new processes and technologies of Co smelting shall be developed to secure the state-of-the-art status of Co smelting and to produce high-end products.

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Abstract:The Pb-Ag residue resulted from hydrometallurgical zinc leaching was studied to recover lead, zinc, carbon, silver, iron and other valuable metals through the combined process of metallurgy-ore beneficiation. The results showed that the direct volatilization yields of lead and zinc were 98.85% and 97.60% respectively when the reductive roasting temperature was 1200℃, the duration was 60minutes, the proportion of coke powder reductant was 40% and the proportion of added lime was 4%; the grades of carbon concentrate, silver concentrate and iron concentrate reached 71.58%, 548.10g/t and 70.55% respectively, and the recovery rates of carbon, silver and iron reached 95.29%, 91.2% and 40.71% respectively by adopting the process comprising carbon roughing-carbon cleaning, silver roughing-silver cleaning, and magnetic separation of iron of the modified roasting slag; and the resource recovery and utilization of lead, zinc, iron, silver in Pb-Ag residue and carbon in harmless-treated tailings was realized through this process.

Abstract:Lead-silver slag contains lead, zinc, silver and other valuable metals, and the treatment process should achieve two purposes: valuable metal recovery and harmlessness. The leached residue from hydrometallurgical treatment process is still hazardous waste; in the pyrometallurgical treatment process, the more advanced Kivcet process and Ausmelt process, together with lead concentrate process to treat lead and silver slag, however, there exist issues like high production and construction costs. This paper introduces a process of co-processing lead-silver slag from zinc smelting and lead concentrate through oxygen bottom blowing smelting. The theoretical calculation shows that if the composition of lead concentrate and lead-silver slag in a smelter is typical, and under the premise that fuel is not required for oxidative smelting, the maximum value of lead-silver slag is 13%. According to practical production data, the lead recovery rate of this process exceeds 98%, the zinc recovery rate exceeds 90%, the silver recovery rate exceeds 98.5%, and the resource recovery rate is remarkable; The coal consumption is 260-330kgce/t, which realizes low-carbon treatment of lead and silver slag; the off gas generated by the process is used for acid production, the dust is returned to the smelting furnace for comprehensive recovery, and the slag can be sold directly to the cement plant for comprehensive utilization, which means the goal of harmless treatment has been achieved.

Abstract:In order to improve the efficiency of zinc hydrometallurgy, this paper draws ε-pH diagram and predominant zone diagram of zinc-containing materials through virtual simulation calculation, which lays a theoretical foundation for zinc hydrometallurgical leaching and roasting. The ε-pH diagrams of zinc sulfide ore and zinc calcine containing Fe, Cu, Pb and Cd at different temperatures and the predominant zone diagrams of Zn-S-O ternary system and Zn-Me-S-O quaternary system of zinc sulfide ore were simulated by using FactSage thermodynamic software. The results show that during direct leaching of zinc sulfide ore, the pH value of Zn²⁺ is higher than those of Fe²⁺, Cu²⁺, Pb²⁺ and Cd²⁺; while during hot acid leaching of zinc calcine, the pH value of Fe²⁺, Fe³⁺ and Cu²⁺ is lower than that of Zn²⁺, the pH value of Pb²⁺ and Cd²⁺ is higher than that of Zn²⁺, but the potential of Zn²⁺ is lower than those of Fe²⁺, Cu²⁺, Pb²⁺ and Cd²⁺; comparing the ε-pH diagrams of two different minerals, it can be seen that the Zn²⁺ stable zone in the leaching of zinc sulfide ore is smaller than that in the leaching of zinc calcine, which indicates that the leaching effect of zinc calcine is better, and indirectly explains the reason why zinc sulfide ore needs to be roasted before treatment. The predominant zone diagram of zinc sulfide ore shows that ZnSO₄ begins to be decomposed into ZnO at the temperature of 873K, and Fe in zinc sulfide ore will combine with Zn to form ZnFe₂O₄, which hinders the formation of ZnO, and ZnFe₂O₄ is not easy to leach, which leads to the decrease of zinc leaching rate. Therefore, the formation of ZnFe₂O₄ should be reduced as much as possible during roasting.

Abstract:Arsenic trioxide is a toxic and harmful substance, and the storage and transportation risks are very high. This paper introduces a hydrometallurgical process for preparing elemental arsenic by using arsenic trioxide, so as to meet the needs of minimization of arsenic trioxide as well as harmless storage and transportation of arsenic trioxide. With arsenic trioxide as a raw material, this process uses R as reducing agent to prepare elemental arsenic through reduction. The reducing agent R in the process can be aluminum powder, cadmium powder or a mixture of these two metal powders in any proportion. The test results show that the reduction rate of arsenic is greater than 99% and the grade of crude arsenic is greater than 97% under the optimum process conditions of reduction temperature being 60℃, reduction time being 5h, initial acidity being 7.5%, and dosage of reducing agent being 1.8times. Elemental arsenic with a grade of more than 99% can be obtained by crude arsenic refining. This process has the advantages of simple process, high recovery rate, low cost and environmental friendliness. It meets the requirements of arsenic trioxide minimization and harmless disposal, and opens up a new way for the commercial application of arsenic solid waste in non-ferrous smelting industry.

Abstract:Aiming at the leaching high-sulfur residue produced in arsenic filter cake treatment process, the elemental sulfur removal process was studied to enrich the valuable metals in the material and realize the recovery and utilization of resources. In this paper, hot leaching method with sulfurizing agent was used to treat the leaching residue of arsenic filter cake. The test results showed that the best leaching conditions were as follows: when the concentration of leaching agent sodium sulfide was 2.00mol/L, the ratio of liquid to solid was 8∶1, and the reaction was carried out at 85℃ for 2h, the leaching rate of elemental sulfur was 98.1% and that of arsenic was 99.0%, while the valuable metals such as bismuth, lead and copper were almost not leached; at the same time, the leachate can be used as sulfurizing agent to precipitate copper ions in waste liquid, and its precipitation efficiency is equivalent to that of sodium sulfide solution, so that the reagent can be fully utilized. This experiment can realize the efficient removal of elemental sulfur in one step and reduce the arsenic content in materials simultaneously.

Abstract:When high arsenic copper concentrate is subjected to pyrometallurgy, arsenic and its compounds mostly volatilize into gas phase (offgas), which leads to a series of problems such as high operation cost of equipment and serious environmental pollution. This paper took the ISA furnace-PS converter-anode furnace intensive copper smelting process as an example, studied the state and migration characteristics of impurity arsenic in ISA furnace smelting, cleaning furnace reduction, converter converting and reverberatory furnace refining, of which 58.8% of arsenic entered ISA furnace dust and 0.4% remained in anode copper. In this test, oxidation volatilization method and alkali addition slagging method were applied, and the control measures of three procedures, namely smelting, converting and refining, were put forward. According to the arsenic characteristics of easy volatility and difficult solidification, the control methods such as injecting alkaline impurity remover and installing ISA furnace guide pipe can improve the arsenic solidification effect and arsenic removal amount in pyrometallurgical stage, and solve the problems of large amount of arsenic-containing dust in smelting and high arsenic content in copper products.

Abstract:The rotary kiln combined with hydrometallurgical process to recover silver from copper anode slime is used by a metallurgical plant. The obtained crude silver powder is of low grade, for which the silver converter process needs to be added, it will result in an increase of energy consumption and severe environmental pollution. In a bid to omit the silver converter process, a new hydrometallurgical process for silver separation is tested by the plant to replace the original silver separation process of ammonia leaching-hydrazine hydrate with the new hydrometallurgical silver separation process. Exploratory tests were conducted on silver separation with sodium sulfite-formaldehyde reduction-oxidative acid leaching process for impurity removal (silver separation with sodium sulfite-formaldehyde direct reduction process) and silver separation with sodium sulfite-decomplexation of silver separation solution-oxidative acid leaching for impurity removal-formaldehyde reduction process(silver separation with sodium sulfite-formaldehyde indirect reduction process). The results showed that the silver separation with sodium sulfite-formaldehyde indirect reduction process featured a sound stability, the grade of obtained crude silver powder reached 99.93% and the content of other impurities approached to zero. If the silver separation with sodium sulfite-formaldehyde indirect reduction process is substituted for the original process, it is estimated that the plant will increase benefits of at least 1386900 CNY. The regeneration and recycling of sodium nitrite solution from the test process is the key issue that needs to be tackled for the follow-up new process. It can not only solve the problem of production water volume expansion and but also lower the volume of wastewater treatment and consumption of sodium sulfite, which are of vital importance to further reduction of production cost.

Abstract:Complex lead-containing samples from comprehensive recovery in pyrometallurgical copper smelting industry contain elements such as Ba, As, Sb, Sn and Bi. In the analysis of lead content, the elements are severely symbiotic and enwrapped due to interaction. The existing detection methods can not solve the problems of complete dissolution of lead and interference of various impurity elements at the same time. In this paper, the acid leaching-alkali fusion method was adopted to treat complex lead-containing samples. The lead leached by acid was measured by EDTA titration, and the lead dissolved by alkali fusion was measured by atomic absorption spectrometry. For acid-leaching solution, sulfuric acid was adopted for separating Ba, hydrogen bromide for removing As, Sb and Sn, and EDTA for Bi complexation, which eliminated the interference of various impurity elements on lead. The method was applied to the measurement of lead in Reference Material Lead Concentrates GBW(E)070173, and the relative error of the obtained result was less than 1%. The method is relatively easy and convenient in operation, with good reproducibility and high accuracy, which is of great reference for accurate measurement of lead in complex lead-containing samples.

Abstract:In this paper, electric arc furnace dust (EAFD), reductant and slag-forming agent were proportioned into carbon-containing pellets and added to the iron bath at 1500℃ in the vertical tube furnace to recover zinc and iron by the iron-bath smelting reduction process. The EAFD and reduced iron slag were subjected to XRF, XRD, SEM-EDS analysis and chemical analysis, and the effects of binary alkalinity R, reduction time and carbon content on dezincification and reductive melting were studied. The results showed that the maximum content of ZnO in the reduced slag was 0.24 wt%, the minimum was 0.07wt%, and the dezincification rate exceeded 99%; when R=1.6, C/O molar ratio was 1.1, and reduction duration was 40min, a majority of the zinc oxides in EAFD were reduced to Zn metal and evaporated into the off-gas, while most of the iron oxides were reduced to metallic iron and entered the iron bath after carburization and the other oxides formed molten slag and built up on the surface of the iron bath after being separated from the molten iron; the increase of the carbon content was conducive to carburization and promoted the slag-iron melting and separation, but hindered the slag-iron separation as the the excess carbon would accumulate in the molten slag when the carbon content exceeds the demand (nC/nO=1.2) for carburization.

Abstract:The rotary kiln boiler dust of a zinc smelter in Xinjiang had a high silicon content, which resulted in the formation of a large amount of colloidal silica in the leachate, causing difficulty in the separation of the solid and liquid. Based on this situation, a combined process of two-stage leaching and recovery of zinc from high-silicon boiler dust + neutralization for silicon removal was developed by taking advantage of the easy leaching feature of soluble silicon under low acid conditions. The results showed that an effective separation of silicon was achieved and the zinc, iron and soluble silicon leaching rates were 61.23%, 3.21% and 95.23% respectively when the endpoint pH value of the first-stage leaching was controlled at 2.0; the silicon removal rate of the first-stage lechate reached above 90% after neutralization by lime and silicon-removal agent for silicon removal; the zinc content in the residue was reduced to 1.43% after high-temperature and high-acid leaching of the first-stage leachate, accomplishing an efficient recovery of zinc in the high-silicon boiler dust; a high-efficiency selective leaching of silicon was realized through the process, the leaching of iron was inhibited and the formation of an unexpected amount of ferric hydroxide colloid was also avoided during lime neutralization for silicon removal.

Abstract:In China, sulfuric acid leaching process is widely used to recover indium from lead and zinc smelting dust. With the increase of tin content in the dust, the leaching rate of indium shows a decreasing trend and is generally less than 60% in sulfuric acid. Targeted at the complex lead-zinc smelting dust containing In₂O₃·xSnO₂, this paper proposes the chlorination leaching process of lead-zinc smelting dust to recover indium. The influence of factors such as sodium chlorate concentration, hydrochloric acid concentration, leaching temperature, sulfuric acid concentration, leaching time, liquid-solid ratio on the leaching rate of indium were investigated. The results showed that the average leaching rate of indium resulted from chlorination leaching reached 92.7% under the following conditions of sodium chlorate concentration of 20g/L, hydrochloric acid concentration of 48g/L, leaching temperature of 80℃, sulfuric acid concentration of 225g/L, leaching time of 3.0h, and liquid-solid ratio of L/S=5∶1(L/kg). Thus, this process can be used in the production practice to recover indium through the treatment of the complex lead-zinc smelting dust.

Abstract:The studies targeted at the separation and recovery of antimony and bismuth were carried out for the Cu anode slime with high-Au & Ag content produced from the electrolysis process of copper smelting enterprises to achieve the purpose of recovering antimony and bismuth from anode slime by adopting chlorination leaching-neutralization & precipitation method, with no precious metals dispersion during this process. Single-factor test was carried out on the factors affecting the leaching efficiency of antimony and bismuth. The test results showed that the leaching rates of antimony and bismuth reached 86% and 98% respectively after cooling and filtration of the solution resulted from 3h reaction under such conditions as the liquid-solid mass ratio of the reaction controlled at 4∶1, the reaction temperature at 80℃, the sulfuric acid concentration of the solution adjusted to 120g/L by using sulfuric acid, and the chloride ion concentration at 140g/L by adding sodium chloride; the unleached copper and arsenic in the copper-removal process were basically leached completely, the leaching rate of lead reached about 10%, and the precious metals gold and silver were further concentrated without being leached. In the test, the antimony and bismuth leachate was subjected to step-by-step neutralization for pH adjustment with relatively high-Sb residue and high-Bi residue obtained under different pH conditions to create conditions for further recovery of antimony and bismuth.

Abstract:Three kinds of refractory bricks, chrome corundum, magnesia-chromium corundum and sintered zirconium corundum, were selected to conduct erosion experiments and to explore the causes of erosion in different acid-base slag environments. The results show that in the acid slag environment, the corrosion resistance of chrome corundum bricks is better, the erosion of magnesia-chromium corundum bricks is more severe, and the erosion of sintered zirconium corundum bricks is obvious. With rising of temperature and time increase, chrome corundum brick comes the best in terms of corrosion resistance. The performance of magnesia-chromium corundum bricks in terms of corrosion resistance and thermal expansion resistance is not good. The erosion of sintered zirconium corundum bricks is more severe. Under neutral slag environment, the corrosion resistance of chrome corundum bricks is reduced, and the erosion is a little bit obvious. The corrosion condition of magnesia-chromium corundum bricks is relatively good, and a certain degree of cracks appear above the liquid surface line. The sintered zirconium corundum brick erodes faster, and there are obvious “grooves” at the liquid surface line. Under basic slag environment, the surface of the chrome corundum brick is eroded and deteriorated obviously. The erosion of magnesia-chromium corundum bricks is not obvious, and the erosion of sintered zirconium corundum bricks is obvious. The analysis on the causes of erosion shows that the raw incinerator slag contains Na₂O amorphous minerals and crystalline quartz, and the added calcium oxides reacts with the chemical composition of the refractory material to cause erosion, and the erosion rate is related to the chemical composition of the refractory material.

Abstract:The As content of arsenic-containing wastewater produced by a smelting enterprise in Shandong is 20251.6mg/L. In this paper, the sulfidization arsenic precipitation process was used to explore better process parameters, and response surface methodology was used for optimization. Test study was made on the effects of sulfidizing agent dosage, reaction time and stirring speed on arsenic precipitation. The optimum technological parameters were obtained as follows: sulfidizing agent dosage n(S^2-)/n(As)=1.2 reaction time 30min, stirring speed 80r/min. Under such conditions, the As content in the post As-precipitation solution was 0.24mg/L. The response surface methodology is used to analyze the significance and interaction of sulfidizing agent dosage, reaction time and stirring speed, and the quadratic regression model obtained is significant, with good fitting degree. The influence sequence of each factor on As precipitation in wastewater is reaction time ＞ sulfidizing agent dosage ＞ stirring speed. The optimum process parameters for response surface prediction were as follows: sulfidizing agent dosage n(S^2-)/n(As)=1.22 , reaction time 38min, stirring speed 64r/min. Under such conditions, the average As concentration in the precipitated solution was 0.15mg/L. The results show that the As concentration in the precipitated arsenic solution is less than 0.2mg/L by using the optimized sulfidizing arsenic precipitation process parameters, which meets the local discharge standard DB37/3416.5—2018 Integrated Wastewater Discharge Standard for Basin Part 5: Shandong Peninsula Basin required by Shandong Province.

Abstract:The isothermal pyrolysis test of waste PCB was carried out by using a self-designed tubular pyrolysis furnace. High-precision electric meter and a special time statistical method were used to improve measurement accuracy and accurately obtain the pyrolysis energy consumption of waste PCB. The test results showed that the measurement accuracy can be further improved by using stopwatch to record the time required for high-precision ammeter to beat 0.01kW·h in pyrolysis process, and make it in the same order of magnitude as the pyrolysis energy consumption of waste PCB so that the error can be reduced to within 3%. The unit pyrolysis heat of waste PCB increased as temperature rose. The unit pyrolysis reaction heat under the optimum pyrolysis temperature of 700℃ was 854.34kJ/kg. And the pyrolysis time gradually decreased with the increase of temperature. The heat value of pyrolysis gas from pyrolysis can fully meet the heat required by PCB pyrolysis reaction. If the capacity of pyrolysis hearth is increased, the waste PCB can be completely self-heated by pyrolysis.

Abstract:The titanium middling ore in the separation product of vanadium titanomagnetite contains 28%-36% TiO2, and the titanium can be enriched from titanium middling ore by the coal-based direct reduction technology. In this paper, the effects of reduction temperature, reduction time, carbon/oxygen molar ratio, raw material particle size and other parameters on the content of TiO2 in non-magnetic products were investigated through coal-based direct reduction experiments. The effects of magnetic field intensity and particle size of non-magnetic products on the content and recovery of TiO2 in products were investigated by magnetic separation experiments. The conclusions drawn are as follows: the optimal process parameters for the coal-based direct reduction process are: binder addition of 3%, molding pressure of 15MPa, reduction temperature of 1350℃, reduction time of 45min, raw ore particle size of 180-200 mesh, carbon ratio nC/nO=1.1; the optimal process parameters of the magnetic separation process are the magnetic field strength of 160mT and the particle size of the reduction product of 180-200 mesh; under this process condition, a non-magnetic product with a TiO2 content of 58.64% and a yield of 63.94% can be obtained. X-ray diffraction was used to analyze the physical images of titanium middles, magnetic products and non-magnetic products. The results show that coal-based direct reduction technology can reduce iron elements in titanium middles into elemental iron, and separate magnetic products and non-magnetic products under the action of magnetic field, so as to realize the separation of titanium iron.