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To recover zinc from electric arc furnace (EAF) dust, a process of primary normal pressure leaching and secondary alkaline pressure leaching is proposed. First, under the alkaline pressure leaching system, the experiment of pure zinc ferrite being reduced by iron powder was carried out. Under the optimal reduction conditions (i.e., temperature of 260 °C, NaOH concentration of 6 mol/L, liquid-to-solid ratio of 50 mL/g, and a 5-fold excess of iron powder), 89% of zinc was extracted. The iron in the reduced residue exists as a magnetite phase. Subsequently, the normal pressure leaching experiment was carried out with EAF dust as raw material, and 66% zinc was leached. The main phase of zinc in normal leaching residue was determined to be zinc ferrite. Then, the normal leaching residue was reduced by iron powder under the alkaline pressure leaching system, and 66.5% of zinc was extracted. After the two-stage leaching process, the leaching rate of zinc in EAF dust can achieve 88.7%. The alkaline pressure leaching solution can be returned as the normal pressure leaching solution, and the magnetite in the alkaline pressure leaching residue can be recovered by magnetic separation.

Background With the application of immune checkpoint inhibitors (ICIs) in cancer treatment, more and more attention has been paid to checkpoint inhibitor-related pneumonitis (CIP), which requires a better understanding of its clinical characteristics and therapeutic effects. Methods The clinical and imaging data of 704 patients with non-small cell lung cancer (NSCLC) who received immunotherapy were analyzed retrospectively; the clinical characteristics of CIP were summarized, and the therapeutic regimens and effects of the patients were summarized. Results 36 CIP patients were included in the research. The most common clinical symptoms were cough, shortness of breath and fever. The CT manifestations were summarized as follows: Organizing pneumonia (OP) in 14 cases (38.9%), nonspecific interstitial pneumonia (NSIP) in 14 cases (38.9%), hypersensitiviy pneumonitis(HP) in 2 cases (6.3%), diffuse alveolar damage in 1 case (3.1%) and atypical imaging manifestations in 5 cases (13.9%). 35 cases received glucocorticoid therapy, 6 patients were treated with gamma globulin and 1 patient was treated with tocilizumab. There were no deaths in CIP G1-2 patients and 7 deaths occured in CIP G3-4 patients. 4 patients were treated again with ICIs. Conclusion We found that glucocorticoid 1–2 mg/kg was effective for most patients with moderate to severe CIP, and a few patients with hormone insensitivity needed early immunosuppressive therapy. A few patients can be rechallenged with ICIs, but CIP recurrence needs to be closely monitored.

Utilizing the abundant biogas resources in an efficient and sustainable way helps to mitigate the energy and environmental problems. CH4 oxy-CO2 reforming reaction is a promising technic and an efficient catalyst for the reaction is in great demand. In this work, Ru was incorporated into monometallic Co/MgO-Al2O3 aerogel catalyst via supercritical drying process combined with sol–gel method and a series of Co-Ru bimetallic aerogel catalyst with different Ru contents was synthesized. Catalytic performance screening for the catalysts indicated that combination of Ru and Co significantly raised the activity of monometallic catalysts, e.g., the activity of Co15-Ru0.2 (Co content: 15 wt.%, Ru content: 0.2 wt.%) was 64.7% and 63.9% higher than those of the monometallic Co15 and Ru0.5 catalysts, respectively. The results also indicated that Co/Ru ratio had a great influence on the catalytic activity: Co15-Ru0.2 > Co15-Ru0.1 ≈ Co15-Ru0.5 > Co15-Ru0.05 > Ru0.5 > Co15. XRD, BET, H2-TPR, XPS, TEM, TG/DSC, and FESEM analyses were performed, disclosing that the reducibility and active metal particle size of the catalysts can be manipulated by the Co/Ru ratio due to the transportation of bulk Co2+ to the surface. The Co15-Ru0.2 catalyst had a mean active metal particle size below 10 nm and medium reducibility, resulting in a high activity and stable performance.

The element partitioning in a Pb−Bi concentrate oxygen-rich bath smelting process was studied using thermodynamic equilibrium simulation method. Effects of oxygen to feed ratio (OFR) and sulfur dioxide partial pressure (pSO2) on the partitionings of Bi, Pb, As, Sb, Cu and Ag were analyzed and compared with industrial data. The results suggested that the optimal OFR was between 6.3 and 6.8 kmol/t to maximize Bi, Pb, Cu and Ag partitioning in the metal phase. Further increase of OFR led to the drop of metal partitioning and increase of slag liquidus temperature. High pSO2 led to high deportment of Bi and Pb in the gas phase mainly in the form of sulfides, suggesting that a low pSO2 was conducive for reducing the dust ratio.

The oxygen-rich bath smelting process of copper dross was simulated using Factsage software. It was found that the two-step simulation results agreed better with the industrial data, especially in the speiss. Bi, Pb, Sb, As, and Cu contents in the speiss by two-step simulation differed from the industrial data by 0.6 wt.%, 0.8 wt.%, 2.6 wt.%, 7.2 wt.%, and 8.6 wt.%, respectively. Two-step simulation results indicated that the speiss partitioning of Sb, As, and Cu increased from 60.1%, 32.3%, and 27.5% to 81.6%, 67.8%, and 47.2%, respectively, with the smelting temperature increasing from 1050 °C to 1250 °C. While the metal partitioning of Sb and Cu increased from 0% and 2.7% to 15.2% and 13.3%, respectively, with the separation temperature increasing from 400 °C to 700 °C.

A complex copper dross containing Bi, Pb, Cu, As and Sb was processed in a 3.6 m2 industrial-scale oxygen-rich side-blown bath smelting furnace (OSBF). The products were collected and analyzed by chemical and mineralogical methods. Industrial statistical data showed that recoveries of Cu, Pb, Bi and Sb reached 87%, 96%, 97% and 60%, respectively. Slag composition indicated that the process can be divided into 2 steps, i.e., oxidative smelting and reductive smelting, due to the use of different oxygen-rich gas flow rate and enrichment. The effect of oxygen partial pressure (PO2) on the element partitioning was simulated by Factsage. The results indicated that increase of PO2 between 10–5 atm and 10–7 atm would significantly increase the partitioning of Cu, Pb, As and Sb to the slag; while element partitioning became almost independent of PO2 between 10–7.5 atm to 10–9.5 atm. Phase analysis for the slag indicated that the bath in the furnace was a “FeO”–CaO–SiO2–Al2O3–ZnO slag coexisting with solid spinels.

The treatment of spent lead paste is essential for the recycling of spent lead-acid batteries. In this study, we propose a facile route for the recovery of lead from spent lead paste by pre-desulfurization followed by low-temperature reduction smelting. The effects of two desulfurization methods, i.e., high-pressure and normal-pressure processes, on the desulfurization efficiency of spent lead paste were compared. The high-pressure process was found to be more effective than the normal-pressure process; 99.1% of sulfur was removed under the high pressure conditions with NaOH excess of 1.1 at 150°C for 30 min. Moreover, a part of PbO2 reacted with PbO to form Pb3O4. Afterwards, the desulfurized residue, mainly consisting of Pb3O4 and PbO2, was directly used in low-temperature reduction smelting, and the optimum smelting conditions were determined to be: starch/Pb-residue mass ratio = 1/6.3; Na2O/Pb-residue mass ratio = 1/15; B2O3/Pb-residue mass ratio = 1/7.5; smelting temperature 700°C; and smelting duration 90 min. The lead recovery under the optimum conditions was as high as 93%. Scanning electron microscope images of the smelting slag showed that the smelting slag mainly consisted of PbOx and Na2O.

To achieve a safe treatment of arsenic-containing acid wastewater, a new process was proposed, including arsenic removal via sulfide precipitation and hydrothermal mineralization stabilization. Under optimal conditions of sulfide precipitation, 99.65% of arsenic from wastewater was precipitated in the form of amorphous As2S3. The As leaching concentration of amorphous As2S3 in TCLP (toxicity characteristic leaching procedure) test was up to 212.97 mg/L, therefore, hydrothermal mineralization was adopted to improve the stability of amorphous As2S3. The results showed that the As leaching concentration of mineralized As2S3 was only 4.82 mg/L. Furthermore, the amorphous As2S3 could be transformed into crystallized As2S3 (orpiment) in the presence of mineralizer Na2SO4. Simultaneously, the As leaching concentration of crystallized As2S3 was further reduced to 3.86 mg/L. Hydrothermal mineralization was an effective method for the stabilization of As2S3. Therefore, this process has a greater application in the treatment of arsenic-containing wastewater.

To overcome the shortcomings of the traditional method, a novel method was proposed to produce sodium pyroantimonate (NaSb(OH)6) with antimony trioxide by pressure oxidation in KOH solution. An E-pH diagram of the Sb-H2O system at different temperatures indicated that various forms of Sb (III) can be easily oxidized to Sb (V) in alkaline solution at high temperature, generating potassium pyroantimonate. The results demonstrated appropriately increasing temperature, oxygen partial pressure, stirring speed, KOH concentration and duration time were beneficial to increasing the Sb leaching ratio. However, a new precipitate could be produced from the leaching solution by excessive temperature. Finally, NaSb(OH)6 should be prepared by adding NaOH in time to avoid the formation of KSbO3 with temperature lowering. The Sb precipitation ratio reached 90.32% under optimal conditions. The results of SEM and particle distribution showed that NaSb(OH)6 is a regular hexahedron with an average particle size of 6.13 µm.

In this article, the methods of lithium salt and potassium salt were used to synthesize LiSb(OH)6. Meanwhile, LiSbF6 was also prepared by the methods of neutralized dissolution and hexafluoroantimonic acid. The relative contents of Li, Sb, and F in LiSb(OH)6 and LiSbF6 were determined, and the results showed these values were extremely similar to the theoretical values. The phases of the products were detected by x-ray diffraction, and the results indicated they were single-phase products with good crystallinity. The morphologies of the products were characterized by scanning electron microscope, which revealed the LiSb(OH)6 crystals were hexagonal layered structures. The infrared spectra of the LiSbF6 crystals proved the presence of SbF6− in the crystals. Moreover, the result of the thermal stability test indicated that the initial decomposition temperature of LiSbF6 was 367.64°C. Therefore, these novel methods can be used as a reference to effectively improve the traditional processes for LiSb(OH)6 and LiSbF6 preparation.

A technical route for preparing sodium pyroantimonate by pressure oxidation in NaOH solution is proposed. The E-pH diagram of the Sb-H2O system shows that Sb(III) from antimony trioxide can be oxidized to Sb(V) to prepare sodium pyroantimonate under different alkaline concentrations. In the direct pressure oxidation technique, the product was doped with antimony trioxide due to the diffusion effect. By comparison, the technique of complex leaching-pressure oxidation could prepare an eligible product, which presented regular hexahedral morphology. Nevertheless, sodium pyroantimonate transformed to NaSbO3 at excessive temperature. The sodium antimonite solution prepared in the leaching process contained 18 g/L Sb. The antimony precipitation ratio in the pressure oxidation process increased with the stirring speed and oxygen partial pressure. Under the optimum conditions of temperature of 150°C, oxygen partial pressure of 2.0 MPa, stirring speed of 1000 rpm, and reaction time of 2 h, the antimony precipitation ratio was 97.70%.

Electric arc furnace (EAF) dust is a hazardous waste material containing valuable elements, such as zinc (Zn) and iron (Fe), as well as other environmentally harmful elements. In the conventional process, the existence of zinc ferrite (ZnFe2O4) in EAF dust results in a low recovery ratio of zinc and can induce high energy consumption or difficulties utilizing iron resources. In this article, a hydrothermal reduction method in a NaOH system is proposed to recover Zn selectively from EAF dust. First, when pure ZnFe2O4 was used as the input material to test the feasibility of the new method, 70% Zn leaching efficiencies and the isolation of Fe in the residue through the formation of ferroferric oxide (Fe3O4) were achieved at the optimal conditions of 260 °C, 6 M [NaOH], 10 g/L [C6H12O6]n and 50 mL/g of L/S. Second, a leaching process of EAF dust was proposed using the following two procedures: (i) first leaching under normal pressure to dissolve zinc oxide (ZnO); and (ii) second leaching of the first leach residue under hydrothermal reduction conditions to dissolve ZnFe2O4. The leach liquor from the second step was recycled to the first step. Under optimum conditions, the leaching efficiencies of Zn during the first and second leaching were 66.4% and 69.3%, respectively. The total Zn leaching efficiency was calculated to be 89.7%, producing Zn-impregnated leach solution and a leach residue rich in Fe (up to 39.4 wt%) that are suitable as resources for a steel-making plant. However, this process also has its shortcomings owning to larger liquid-solid ratio in the leaching process. Overall, the study technically provides an alternative approach to improve the recycling of zinc and iron and effectively reduce environmental risks from EAF dust.

Co–Pt/MgO–Al 2 O 3 bimetallic aerogel catalysts were synthesized via a sol-gel combined with supercritical drying method. The catalysts were characterized by XRD, BET, HRTEM, STEM-HAADF, XPS, H 2 -TPR, H 2 -TPD, TG/DSC, FESEM and their catalytic performances in CH 4 oxidative CO 2 reforming were evaluated. The H 2 spillover effect between Pt and Co enhanced the reducibility of the catalyst, while the strong metal-support interaction (SMSI) effect in the bimetallic aerogel catalysts confined the agglomeration of metal particles. Pt/Co ratio played a key role on the existence of surface metal species, leading to different catalytic performances. The optimal Pt/Co ratio was Pt/Co = 0.02 w/w, on which a 50% higher activity in terms of CH 4 conversion than monometallic Co or Pt aerogel catalysts was obtained. Whereas the impregnated catalyst with an identical composition showed a much lower activity. The Co–Pt aerogel catalysts also showed high resistance to inactive carbon formation. The oxidation temperature of the carbon species deposited on the spent Co–Pt aerogel catalyst was only 275 °C and no filamentous or graphitic carbon was identified, disclosing that the formation of inactive carbon was inhibited due to the synergy between Co and Pt and the SMSI effect.

CH4 temperature programmed surface reaction (TPSR) and CH4 oxy-CO2 TPSR are applied to test Ni (10Ni), Co (10Co) and Co–Ni (5Co5Ni) aerogel catalysts in the CH4 oxy-CO2 reforming reaction. Monometallic 10Ni outperforms 10Co and bimetallic 5Co5Ni aerogel catalysts in the CH4-TPSR test. The CH4 activation temperature (AT) of 10Ni and 5Co5Ni is 160 °C lower than that of 10Co, indicating that CH4 can be activated more effectively on 10Ni and 5Co5Ni. As a result, the critical temperature (CT) of 10Ni and 5Co5Ni in the CH4 oxy-CO2 TPSR test are 680 °C and 695 °C, whereas 10Co shows no CT in the test. The CH4 oxy-CO2 TPSR test also reveals that although CTs of the catalyst are independent of the reduction pretreatment, the catalytic oxidation abilities for carbon species are greatly promoted after reduction. The calcined samples activate CH4 via the combination of combustion and partial oxidation, while the reduced sample activate CH4 mainly via combustion. Furthermore, the reduced 5Co5Ni activates and completes CH4 oxidation at temperatures 60 °C and 128 °C lower than those of 10Ni, suggesting its higher ability to catalyze the oxidation of carbon species generated in CH4 reforming, which could be account for its better performance than those of the monometallic catalysts.

To study the effect of temperature on the stability constant of zinc-nitrilotriacetic acid- ammonia (Zn-NTA-NH3) system, the protonation constants of the free ligands and stability constants of the ternary complexes were determined via pH potentiometric method at temperatures of 25.0 °C, 40.0 °C, 55.0 °C, 70.0 °C. The first and second stability constants of Zn-NTA-NH3 system decrease from 11.83, 14.00 to 10.72, 12.81, respectively, with the temperature increasing from 25.0 °C to 70.0 °C. While those of the Zn-NTA-H2O system decrease from 9.20, 11.98 to 8.27, 11.29. These results indicate that Zn-NTA-NH3 system is more stable than Zn-NTA-H2O system, leading to enhanced zinc leaching efficiency in the zinc leaching process.

An innovative process was proposed for recovering lead from spent lead paste, and it produced less pollution and used less energy than the traditional process. First, lead dioxide in lead paste was reduced by glucose under hydrothermal conditions. The effects of the reaction time, glucose excess coefficient, temperature and pH on the hydrothermal reduction were systematically investigated. Under the optimized reduction conditions (i.e., temperature of 175 °C, time of 120 min, glucose excess coefficient of 3.0 and pH of 5.5), 99.9% reduction ratio of lead dioxide is achieved, and only the PbO·PbSO4 and PbSO4 phases are observed in the reducing residue. Subsequently, the reducing residue is desulfurized in a NaOH solution, and approximately 99.40% of the sulfur is removed. The main lead phase in the desulfurization residue is 3PbO·H2O.

In view of the difficulties in separating similar metals from acidic wastewater, a new process is proposed to selectively separate metals by controlling potential. In this study, the effects of solution potential on precipitation ratio and separation efficiency were investigated. The potential was controlled by adding sodium sulfide or xanthate and was monitored by an MT320-SpH instrument with a Pt electrode and a calomel electrode. E-pH diagrams of Me-S-H2O system were plotted through thermodynamic calculations. The results revealed that with the drop of solution potential, the metal precipitation ratios increased. Cu2+, Co2+, and Zn2+ can be selectively separated and completely precipitated under controlling potential at 330 mV, 125 mV and 30 mV, respectively. Mn2+ could be recovered by carbonate precipitation from the residual solution after Zn removal. The precipitation ratios of Cu2+, Co2+, Zn2+, and Mn2+ were as high as 99%.

In the face of difficulties in separating similar metals from acidic wastewater, a new process is proposed to selectively separate metals from acidic wastewater by controlling potential. In this paper, the effect of pH value on precipitation ratio and separation efficiency is investigated. Through thermodynamic calculations, a 3D E -pH diagram of the Me-S-H 2 O system was plotted. The results reveal that rising pH value is helpful for metal precipitation, but high pH value may cause co-precipitation of metals impurity. The crystallinity and size of precipitates are obviously affected by pH value. Co can be effectively separated from Zn by xanthate precipitation under controlling potential, and the xanthate dosage is dependent on pH value. Mn-inclusions appear in Zn-precipitate due to high Mn 2+ concentration. The optimal pH value for Cu 2+ , Co 2+ , and Zn 2+ precipitation were 1, 2, and 2, respectively. Mn 2+ could be recovered by carbonate precipitation from the solution after Zn removal. Cu 2+ , Co 2+ , Zn 2+ , and Mn 2+ precipitation ratios reached to 99%.

A study was conducted for metal extraction from copper smelting dust using the oxidation leaching and control of potential technology. The effects of H2O2 dosage, H2O2 feeding speed, initial HCl concentration, leaching temperature, liquid-to-solid ratio and leaching time on metals leaching efficiencies were investigated. The following optimized leaching conditions were obtained: H2O2 dosage of 0.8 mL/g (redox potential of 429 mV), H2O2 feeding speed of 1.0 mL/min, initial H2SO4 concentration of 1.0 mol/L, initial HCl concentration of 1.0 mol/L, leaching temperature of 80 °C, initial liquid-to-solid ratio of 5:1 mL/g and leaching time of 1.5 h. Under the optimized conditions, copper and arsenic can be effectively leached from copper smelting dust, leaving residue as a suitable lead resource. The average leaching efficiencies of copper, arsenic and iron are 95.27%, 96.82% and 46.65%, respectively.

A novel process was proposed for the separation of Cu and As from Cu-smelting waste acid (CSWA). This process includes three steps, namely, the oxidization of As via the addition of H2O2, the selective separation of Cu through the use of Na2S, and the removal of As via the synthesis of scorodite. The CSWA was initially oxidized using H2O2 where the As(III) transformed into As(V). The oxidization of As(III) is believed to be the decisive step with regard to the selective separation of Cu and As via the sulfide precipitation process. The results showed that 99.4% of the Cu and 1.8% of the As were precipitated via the sulfidization process and that 98.8% of the As in the Cu-removed CSWA was removed in the form of scorodite (FeAsO4·2H2O). This novel process exhibits promising application prospects for the treatment of CSWA.

Antimony-bearing dust, a byproduct of the lead-smelting process, is an important secondary resource of antimony. A clean production process for producing NaSb(OH)6 by oxidation in an NaOH solution has been proposed to solve the difficulties of recovering antimony from antimony-bearing dust. First, antimony trioxide in the dust was leached into a sodium hydroxide solution. Then, the lead impurities were removed from the leaching solution with the addition of Na2S. Finally, sodium pyroantimonate was prepared by ambient or pressure oxidation from the lead removal solution. Furthermore, the oxidized solution was reused in the leaching process. This study shows that the leaching recoveries of antimony, arsenic, and lead in the first step were 75.04%, 99.18%, and 37.78%, respectively. The lead removal efficiency was 94.04% for the leaching solution in the second step. Under ambient and pressure oxidation, the precipitation rates of antimony were 95.82% and 98.76%, respectively, and arsenic was not detected in the product. The Sb3+ content in the synthesized sodium pyroantimonate product was less than 0.1%, and the product exhibited a regular tetragonal morphology.

Conventional cyanidation technology achieves low gold recovery when used to process refractory gold concentrate. Based on the geochemical characteristics of gold deposit mineralization, a new method is proposed herein for gold enrichment in antimony matte by smelting of refractory gold concentrate. The effects of the FeO/SiO2 and CaO/SiO2 ratios, smelting temperature, and smelting time on the gold recovery were investigated in detail. The optimum conditions were determined to be FeO/SiO2 ratio of 1.2, CaO/SiO2 ratio of 0.4, smelting temperature of 1200°C, and smelting time of 45 min. The gold content in antimony matte and smelting slag was 96.68 and 1.13 g/t, respectively. The gold, antimony, and arsenic recovery was 97.72%, 26.89%, and 6.56%, respectively, with most of the antimony and arsenic volatilized into dust. Mineral liberation analyzer results showed that the antimony matte mainly consisted of FeS and FeO, with three phases, viz. FeAs, SbAs, and AuSb, embedded between them, indicating that gold was easily enriched with antimony and arsenic during smelting of refractory gold concentrate.

Oxygen-rich side-blow bath smelting (OSBS) technology offers an efficient method for processing complex bismuth-lead concentrates; however, the element distributions in the process remain unclear. This work determined the distributions of elements, i.e., bismuth, lead, silver, copper, arsenic and antimony, in an industrial-scale OSBS process. The feed, oxidized slag and final products were collected from the respective sampling points and analyzed. For the oxidative smelting process, ~ 65% of bismuth and ~ 76% of silver in the concentrate report to the metal alloy, whereas less lead reports to the metal (~ 31%) than the oxidized slag (~ 44%). Approximately 50% of copper enters the matte, while more than 63% of arsenic and antimony report to the slag. For the reductive smelting process, less than 4.5% of bismuth, lead, silver and copper in the oxidized slag enter the reduced slag, indicating high recoveries of these metal values.

The Sm Co alloy waste is considered to be one of the most valuable wastes for recycling, primarily due to the value of the Co and Sm metals. The two metals were extracted from the Sm Co alloy waste by sulfuric acid leaching, and Sm(III) was separated by sodium sulfate as a precipitate of sodium samarium sulfate. It was found that 99.9% of the Co and 24.1% of the Sm were extracted into the leach solution under the most suitable conditions: liquid-solid ratio of 6:1, leaching temperature of 60 °C, leaching time of 75 min and H2SO4 concentration of 4 mol/L. A high fraction of 93.4% Sm was precipitated at 80 °C, and the Na2SO4 to Sm molar ratio was 4:1 with 60 min of precipitation time. Thus, sulfuric acid leaching followed by sodium sulfate precipitation is feasible for recovering Co and Sm from the Sm Co alloy waste.

The pressure oxidation process to prepare sodium pyroantimonate from sodium thioantimonite solution was optimized. The results indicate that increasing reaction temperature, oxygen partial pressure and stirring speed can significantly accelerate the oxidation precipitation rate of Sb. A decrease in Sb3 + concentration is beneficial for shortening the reaction time. The NaOH concentration has no significant effect on the oxidation velocity but plays an important role in the quality of sodium pyroantimonate. An increase in stirring speed and oxygen partial pressure result in the formation of a red precipitation in the solution with 12.5 g/L of NaOH, while the desired sodium pyroantimonate can be produced with a NaOH concentration of 30.0 g/L. The XRD and SEM-EDS analysis confirm that the red precipitate is composed of NaSb(OH)6 and Na6Sb4S9.

In this study, we report the recovery of metals from copper smelting dust by oxidation leaching and fractional precipitation technology. Under the optimum conditions, the leaching efficiencies achieved for Cu, As, Fe, Cd, and Zn are 93.4%, 94.2%, 39.7%, 98.1%, and 90.7%, respectively. Fractional precipitation technology is then employed for the recovery of Cu, As, Zn, and Cd from the leaching solution. The precipitates generated in the sequential steps contain 54.2% Cu, 25.8% As, 65.3% Cd, and 59.5% Zn. The precipitation of Cu, Cd, and Zn as their sulfides can be modulated by controlling the redox potential, and the optimal values are +312 mV, +126 mV, and +30 mV, respectively.

In this work, the leaching behavior of the major components, i.e., Zn, Pb and Fe, of refractory hemimorphite in an ammonium chloride-ammonia-nitrilotriacetic acid (NH 4 Cl-NH 3 -NTA) system was investigated in detail. The thermodynamic analysis of the Zn 2 + -NTA 3- -Cl - -NH 3 -H 2 O and Pb 2 + -NTA 3- -Cl - -NH 3 -H 2 O systems demonstrated that the predominant zinc and lead species in the leaching solution were the ZnNTA(NH 3 ) 2 - and PbNTA - complexes, respectively. The results of the leaching experiments indicated that more than 85% of Zn, 60% of Pb and less than 3% of Fe were leached from refractory hemimorphite using 2 mol/L NH 3 –2 mol/L NH 4 Cl–0.15 mol/L NTA as the leaching agent at 40 °C with a liquid-solid ratio of 10 mL/g. Moreover, a feasible method for zinc recovery and ligand regeneration was proposed: the ZnNTA(NH 3 ) 2 - complex leaching solution was transformed into a zinc chloride solution through ammonia distillation and an acidity adjustment process. More than 96.0% of the NTA was recovered when the leaching solution pH was reduced to 1.0.

The nonferrous metals in blast furnace dust, especially zinc, which used to be categorized as hazardous components for blast furnace, are worthy of recovery in this economy. In this paper, a novel processing for extracting zinc from blast furnace dust by using a coordination reaction between the organic ligand and zinc ions was proposed. When blast furnace dust was leached in a 0.2 M iminodiacetic acid (Ida) solution at 20 °C for 2 h with a liquid-to-solid ratio of 10 mL/g, the leaching rates of zinc and iron were 62.78% and 6.07%, respectively. To accomplish the maximum separation of zinc from iron during the leaching process, a mixed-ligand complex leaching system consisting of Ida as the primary ligand and ammonium as the secondary ligand was developed. The optimal conditions for the leaching process with a mixed-ligand complex leaching system were investigated. We found that 65.58% of zinc was extracted by the solution with Ida:NH 4 Cl:NH 3 ·H 2 O at 0.2:2:2 M and a 20:1 liquid-to-solid ratio at 40 °C for 2 h.

To avoid the adverse effects of antimony on the cyanide leaching process used for the recovery of gold from refractory gold ore, a cleaner production process was proposed that would selectively remove antimony as sodium pyroantimonate. The process included three steps: 1) leaching of antimony from the refractory gold ore with sodium sulfide solution; 2) pressure oxidation of the leaching solution containing sodium thioantimonite, transforming it to sodium pyroantimonate; and 3) concentration/crystallization to recover the sodium thiosulfate byproduct from the oxidized solution. The leaching recoveries of antimony, gold, and arsenic in the first step were 96.64%, 1.44%, and 0.41%, respectively. The precipitation ratio of antimony exceeded 99.80% during the second step. The synthesized sodium pyroantimonate product exhibited a regular tetragonal morphology. The purity of the sodium thiosulfate byproduct obtained during the third step reached 98.0%.

An environmentally friendly process for the selective extraction of zinc from refractory hemimorphite using iminodiacetic acid (H2Ida) as a complexing agent was investigated. Thermodynamic simulations demonstrated that the partially dissociated protons and amino carboxylate anions synergistically affect the dissolution of hemimorphite. The experimental results were consistent with the theoretical analyses. Under optimal conditions, the leaching extraction of zinc was above 88% and that of iron was below 4%. Furthermore, the complexing agent could be recovered from the leaching solution at the isoelectric point by adding dilute sulfuric acid.

Ni-MgO/Al 2 O 3 (Ni-A) and Co-MgO/Al 2 O 3 (Co-A) aerogel catalysts are mechanically mixed with different weight ratios and their catalytic performances in CH 4 reforming with CO 2 and O 2 are investigated. The catalyst mixtures showed higher CH 4 conversions than those theoretical values predicted from the mixture law, indicating an interesting promotion effect between them. The catalysts are characterized by XRD, BET, SEM, TG/DTG and Raman spectrum techniques. It is disclosed that the reduced Co-A can be easily oxidized during the reaction, but Ni-A can prevent its oxidation during the reaction, which enables Co-A in the mixture to maintain its catalytic reforming ability. CH 4 conversion is raised dramatically for ∼50% as comparing the catalysts mixtures with Co-A. On the other hand, the amount of carbon can be decreased by ∼17 wt.% with the increase of Co-A’s weight ratio in the mixture. Meanwhile, carbon species on the spent catalysts turn from inactive graphite to active carbon species. This could be attributed to the strong affinity of Co-A for oxygen species and the decreased active metal particle size in the mixtures.

A hydrothermal method was adopted to strengthen the desulfurization reaction of PbSO4 with the aim of addressing the problem of incomplete desulfurization and large reagent consumption in the pretreatment process of spent lead paste. Mineral liberation analysis showed that lead sulfate disseminates in lead oxide and they become mixed with each other, which is an important reason for incomplete desulfurization via conventional methods. E-pH diagrams of the Pb-SO42-–H2O system at different temperatures indicate that the pH value of PbSO4 converted to PbO decreased with an increase in temperature. Experimental results showed that increases in time, the NaOH excess coefficient, temperature, stirring rate, and the liquid-solid ratio are beneficial for improving the desulfurization ratio. Under hydrothermal conditions, PbO and PbO2 react to form Pb3O4, and this comproportionation reaction destroys the encapsulation of the solid product layer and is conducive to the progress of the desulfurization reaction. Under the optimized reduction conditions (i.e., time of 60 min, sodium hydroxide coefficient of 1.1, temperature of 150 °C, liquid-solid ratio of 10 mL/g, and stirring rate of 600 rpm), a 98.95% desulfurization ratio was achieved. The phases of lead in the desulfurization residue were mainly PbO, PbO2, and Pb3O4, and these can be recovered via low-temperature reduction smelting.

In this study, a novel, efficient, and eco-friendly method is proposed to precipitate Cl− as BiOCl from MnSO4 electrolyte using (BiO)2SO4 as the chloride removal agent. The effects of different factors on the transformation and activation of Cl− removal agent, chloride precipitation, recovery and recycling, and purification processes were investigated. Increasing the L/S ratio could promote transformation and activation, while low temperature and low pH could have positive effects on Cl− precipitation. Under optimal conditions, the chloride removal efficiency reached 95.25% and the concentration of Cl− remaining in the electrolyte was reduced to 43.96 mg/L. Following the recovery and recycling process, the recycled agent (BiO)2SO4 worked well for Cl− precipitation; subsequently, chloride open this circuit as HCl. In purification process, manganese powder was used to reduce the concentration of Bi3+ that remained in the electrolyte to avoid the introduction of new elements. Thus, the concentration of Bi3+ was reduced from 125.00 mg/L to 1.80 mg/L, and the reduced bismuth could be returned to the transformation and activation of Cl− removal agent process. This entire process involving the main elements in the reactions followed a repeatable procedure, as shown in the schematic, wherein 99.98% of bismuth could be recycled and 95.25% of chloride opened the circuit in the form of HCl.

Carbonate decomposition of carbonic refractory gold ore and the following pressure oxidation were studied. In the carbonate decomposition procedure, the effects of liquid-to-solid ratio and reaction time on decomposition ratio of carbonate were investigated. The experimental result shows that the decomposition ratio of carbonate is 98.24% under the conditions of liquid-to-solid ratio of 5:1, Fe3+ concentration of 20 g/L, sulfuric acid concentration of 20 g/L, reaction temperature of 80 °C and reaction time of 2 h. Then, the slurry obtained from carbonate decomposition was put into the titanium autoclave for pressure oxidation leaching. Effects of liquid-to-solid ratio, temperature, time and oxygen partial pressure on sulfur oxidation ratio were studied during pressure oxidation. With the prolonged time, pyrite and arsenopyrite are oxidized to ferric subsulfate, hydrated ferric sulfate and jarosite, resulting in the increasing residue ratio. The residue ratio and the sulfur content in the residue can be decreased by ferric subsulfate dissolution. The oxidation ratio of the sulfur is 99.35% under the conditions of oxidation time of 4 h, temperature of 210 °C, oxygen partial pressure of 0.8 MPa and stirring speed of 600 r/min.

Reduction smelting of the bismuth oxide residue from pressure leaching of bismuth sulfide was investigated in the FeO-SiO2-CaO ternary slag system. The results show that all the recovery ratios of Bi, Ag, Cu and Pb increase with the increase of reductive coal proportion, reaction temperature and time, while too much reductive coal would help Fe enter metal phase; CaO/SiO2 and FeO/SiO2 of the chosen slag system should be 0.5-0.75 and 1.25-1.75, respectively, for the reason that the slag system has the optimum mobility and is beneficial for the recovery of metals. The corresponding optimum conditions are determined as follows: the added coal proportion is 7% of the leaching residue, CaO/SiO2 mass ratio in the chosen slag system is 0.5 and FeO/SiO2 is 1.5, the reaction temperature is 1300 °C and the reaction time is 40 min. Under the above conditions, the recovery ratios of Bi, Ag, Cu and Pb are 99.6%, 99.8%, 97.0% and 97.3%, respectively.

A novel method for leaching of low grade zinc oxide ores in NH 4 Cl–NH 3 solutions containing nitrilotriacetic acid (NTA) as complexing agents is proposed in this paper. The influences of NTA concentration, leaching temperature, stirring speed, particle size and liquid–solid ratio on the zinc extraction are discussed. It is found that the addition of NTA facilitates the zinc extraction. Under the condition that particle size is 64–71 μm, NTA concentration is 0.1 mol/L, liquid–solid ratio is 20 mL/g, leaching temperature is 40 °C and stirring speed is 400 rpm, the zinc extraction reaches 90.3%. The result of leaching kinetic study indicates that the leaching process is controlled by the interface transfer and diffusion through the product layer and the corresponding activation energy is 22.3 kJ/mol.

Copper anode slime is an important source for extracting tellurium. In this study, a novel process for extracting tellurium from the calcine of copper anode slime via continuous enrichment is proposed. First, tellurium in the calcine of copper anode slime is enriched in the replacement residue by oxidative leaching and replacement with copper powder by controlling potential. Second, tellurium in the replacement residue gets further enriched in tellurium concentrate by sulfation roasting and sulfuric acid leaching. Finally, tellurium ingot is prepared from the tellurium concentrate by alkaline leaching and electrowinning. The results indicate that increasing the concentration of hydrochloric acid is beneficial to the leaching of tellurium and copper during the oxidative leaching process, and the leaching rates of tellurium and copper are 72.86% and 99.25%, respectively. In the replacement process, 99.5% of tellurium is replaced when the solution potential is controlled at 279 mV, and tellurium in the replacement residue exists as cuprous telluride. When the calcine of replacement residue is leached by sulfuric acid, the leaching rate of copper is 98.25%, while that of tellurium is only 3.0%. The tellurium content of the tellurium concentrate is 35.58%. In the alkaline leaching of tellurium concentrate, the leaching rate of tellurium is 97.5%. Tellurium ingot, with a purity of 99.99%, is produced from leaching solution by purification and electrowinning. The purity reaches the No.1 standard in YS/T222-1996 of China. The process achieves efficient extraction of tellurium from copper anode slime by continuous enrichment.

Spent lead paste, an important secondary resource, calls for recycling at the request of both environmental protection and the sustainable development of lead industry. An environment-friendly process to recover lead from spent lead paste was proposed. The impurities Sb, Ca, Fe and Cu in spent lead paste were firstly removed by H2SO4 at the leaching rate of 92.22%, 86.43%, 87.89% and 94.04%, respectively. And then, 99.9% of PbO2 in spent lead paste was reduced to PbO by the hydrothermal reduction process with glucose as the reductant. In the process of hydrothermal desulfurization by NaOH, 99.46% of S in the reduced lead paste was removed and then the lead paste was converted into a single phase of 3PbO·H2O. Finally, metallic lead, with a purity of 99.978%, was obtained from the single-phase lead paste by a low-temperature reduction smelting process with starch as the reductant under the conditions of Na2O/B2O3 = 1/2, PbO/B2O3 = 5/1, starch/PbO = 1/7.5, smelting temperature 700 °C, and smelting duration 90 min, and the lead recovery is high up to 95.53%. Compared with the traditional lead smelting technology, the smelting temperature was reduced from 1200 °C to 700 °C, helping greatly reduce the energy consumption. The process proposed in this paper guides a new development direction for the utilization of spent lead paste.