Recovery Of Bismuth Research Articles

Comparison of the electrospun nanofiber discs and polymer inclusion membranes composed of PVC/Aliquat® 336 for the extraction of Bi(III) from zinc ingots

The abilities of PVC-based electrospun nanofiber discs (ENDs) and polymer inclusion membranes (PIMs) containing the extractant Aliquat® 336 for the recovery of Bi(III) were appraised. The ENDs were developed by using a mixture of PVC/Aliquat® 336 (70/30 wt%) dissolved in a mixture of THF/DMF (60/40 v/v%). The investigated PIMs were prepared by using the same composition via the casting method, while THF was used as the solvent. It was observed that Bi(III) (initial concentration 20 mg L−1) from 50 mL solution of HCl/H2SO4 (0.2/0.1 mol L−1) is quantitatively extracted by the selected ENDs and PIMs (3.5 cm in diameter). Regardless of the discs type, the back-extraction was performed by 0.2 mol L−1 EDTA/NH3. Some approved advantages of the ENDs over the PIMs were: (1) the kinetics of extraction and back-extraction process by the ENDs was significantly faster than that observed for the investigated PIMs, (2) the specific surface area of the ENDs was ∼5 times greater than the PIMs, (3) the calculated maximum capacity of ENDs was about 10.5 times greater than that of the PIMs, and (4) the reusability of ENDs was slightly better than PIMs. It was confirmed that the extraction of Bi(III) is taken place by formation of the anionic chloro-complexes BiCl4-, via an anion exchange mechanism. The applicability of the studied ENDs and PIMs was examined by their use in the recovery of bismuth from zinc ingots industrial sludges.

A novel process for extracting bismuth from high iron content copper smelting dust by magnetic separation and leaching process

In this study, an efficient method for extracting bismuth from copper smelter flue dust with high content of iron was proposed. In this method recovery of bismuth from copper flue dust was carried out after copper and iron separation. To separate copper from dust, this material was leached with water and dilute sulfuric acid (0.6 M). Magnetic separation was performed with field intensities of 0.1 T and 0.4 T to separate iron in the leaching residue. 70 % of bismuth is placed in the non-magnetic fraction. The non-magnetic fraction of this separation was leached with (1.4 M) H2SO4 + (1 M) NaCl solution to recover bismuth. In optimal conditions, more than 98 % of bismuth in the non-magnetic fraction is leached. Bismuth precipitation from this solution was done by adjusting the pH to 2.5 with the ammonia solution. Impurities such as; Pb, Fe, As, and Cu were precipitated with bismuth in this condition. Lead was removed as a PbCl2 compound by leaching the precipitate with 6 M hydrochloric acid. Other elements were leached in this chloride environment and entered the solution phase. Bismuth was separated from this leaching solution by hydrolysis in the form of BiOCl. The purity of the BiOCl product was about 95 %. This substance can be used in photocatalyst and solar cell industries.

Thermodynamic and kinetic analyses and experimental verification of tellurium and bismuth recovery from bismuth telluride waste by selective sulfidation-vacuum volatilization

The recovery of tellurium and bismuth from bismuth telluride waste is of great significance in improving the utilization levels of renewable resources, protecting the ecological environment and alleviating the problem of resource shortages. In this paper, tellurium and bismuth were recovered from bismuth telluride waste by selective sulfidation–vacuum volatilization. The Gibbs free energy, phase diagram and thermodynamic equilibrium diagram for the core reaction were calculated and analyzed to determine the possible reactions and the conditions required for the stability of bismuth telluride during the sulfidation process. After calculating the saturated vapor pressures, it was concluded that tellurium and bismuth sulfide could be separated by vacuum volatilization. The thermogravimetric analyses and differential scanning calorimetry were then used to analyze the reaction between bismuth telluride and sulfur. The results indicated that the temperature range for the sulfidation reaction should be 491–620 K. The volatilization temperature of tellurium should be greater than 695 K. The evaporation kinetics for tellurium and bismuth sulfide showed that the evaporation rate of tellurium was much higher than that of bismuth sulfide. Furthermore, experiments were conducted to verify the reliability of the theoretical analysis, and these were in good agreement with the thermodynamic and kinetic results. The direct yields of tellurium and bismuth were 99.65% and 99.49%, respectively. These results confirmed that the selective sulfidation-vacuum volatilization method is a feasible alternative for recovery of tellurium and bismuth from bismuth telluride waste.

Recovery of bismuth and other metals from blast furnace dust by leaching with oxalic acid-based deep eutectic solvent and precipitation

Eutectic solvents, with the advantages of easy synthesis, high chloride content, non-aqueous environment, and low costs are the most promising alternatives to ionic liquids for leaching and recovery of metals from waste resources. The leaching of bismuth has been mainly conducted using chlorinated salts and strong acids in previous studies. The use of conventional ionic liquids can be costly. In addition, there has only been limited research into developing strategies for the recovery of bismuth from blast furnace (BF) dust using a deep eutectic solvent (DES) made of choline chloride and oxalic acid. In this work, an oxalic acid-based deep eutectic solvent was used for selective recovery of bismuth from blast furnace dust. A leaching efficiency of 94.9% Bi was achieved at the most suitable conditions (stirring speed: 300 rpm, liquid-solid ratio: 5 mL/g, temperature: 70°C, time: 20 h). The zinc and iron which are impurities in the leach solution were removed by hydrolysis and photoinduced reduction process, and the by-products of ZnC2O4·2H2O and FeC2O4·2H2O had purities of 86.6% and 96.8%, respectively. Another hydrolysis was employed to recover Bi(III) from the leach solution and the product of BiOCl was obtained with a total Bi recovery efficiency of 89.1%. The proposed process in this work is resource-saving and friendly to the environment.

Selective sulfidation-vacuum volatilization processes for tellurium and bismuth recovery from bismuth telluride waste thermoelectric material

Bismuth telluride-based alloy materials are currently the best performing thermoelectric materials at near room temperature; however, their production and use generate waste (e.g., cutting waste and failed grains). There is also lack of efficient recycling strategies for the generated waste. In this study, a selective sulfidation-vacuum volatilization method is proposed for recovering bismuth telluride waste. The Gibbs free energies of the sulfidation reaction of bismuth telluride are calculated, the saturated vapor pressure of each substance is analyzed, and the composition of the products is predicted. Based on the differences among the sulfidation and volatile properties of bismuth and tellurium, by adding sulfur to bismuth telluride waste, the composition of the substances was regulated, and efficient separation of tellurium and bismuth was achieved. We combined theoretical calculations and experimental studies to investigate the effect of process conditions on the separation and recovery of tellurium and bismuth. The results show that bismuth was thoroughly sulfereted and tellurium was a pure metal when the mass ratio of sulfur to bismuth telluride was 0.168, the sulfidation temperature was 573 K, and the holding time was 60 min. After sulfidation of the bismuth telluride waste, the sulfides were telluride and bismuthous sulfide. The sulfides, that resulted from sulfureted bismuth telluride production, were treated via vacuum volatilization. The optimal vacuum volatilization condition was 873 K for 120 min. The purities of tellurium and bismuth sulfide obtained by the selective sulfidation-vacuum volatilization experiment were >99%. The distribution ratios of tellurium and bismuth were 98.46% and 99.59%, respectively. The method thoroughly separated tellurium and bismuth from bismuth telluride waste, considerably reducing the environmental and economic costs compared with those of the conventional processes.

Green and short smelting process of bismuth sulphide concentrate with pyrite cinder

The traditional smelting process of bismuth sulphide concentrate faces problems of low metallic bismuth recovery, high production cost and low-concentration SO2 flue gas pollution. In this study, a clean and efficient smelting process was developed to directly extract bismuth metal from bismuthinite concentrate. The reaction mechanism and phase transformation laws were confirmed by thermodynamic simulation calculation, TG-MS thermal analysis and phase characterization. Firstly, Fe2O3 and Fe3O4 interacted with Bi2S3 to form Bi2O3. The intermediate Bi2O3 was then reduced into metallic Bi, and the element S was fixed as FeS. Results of the smelting simulation experiments showed that 89.25% metallic Bi was recovered. About 38.17% Pb, 33.92% Cu, and 83.98% Ag were enriched in crude bismuth, and 96.37% S was fixed in the matte as raw material for resource recovery and energy utilisation.

New Processes to Treat Anode Slime and Other Smelter Improvements

During efforts to solve common technical challenges to the comprehensive recovery of rare precious metals associated with copper smelting, Xiangguang Copper (XG) independently developed a new process to treat anode slime using technology that captures precious metal using bismuth. This paper focuses on the introduction and application of this technology for capturing precious metals with bismuth. It also discusses silver electrolysis at high current density and bismuth recovery from precious metals slag. The new process shows good adaptability of raw materials, elimination of lead pollution, low energy consumption, environmental protection and high comprehensive recovery rate of metals, which promotes the technological progress of comprehensive recovery and utilization of associated resources in copper smelting process.

Electrolytic Processing of Pb-Bi Alloy

Metal bismuth is mainly produced as a by-product in the production of lead, tungsten, copper, silver, gold, tin and zinc. Approximately 90 % of all extracted bismuth is obtained from lead, copper and other concentrates. The main source of bismuth is lead concentrates obtained during the processing of lead, as well as lead-zinc and other polymetallic ores. During the processing of these concentrates, bismuth almost completely enters the rough lead, from which it is removed during its refining. The most common technologies for the recovery of bismuth from lead ingots are the Kroll-Betterton process and the Betts electrolytic process. During the electrolysis of the Bi-Pb alloy, the separation of three products has been established, they are anode and cathode alloys, as well as salt melt. The complexity of pyroelectrometallurgical processing of a bismuth-poor alloy with the production of rough bismuth in one stage is confirmed, which necessitates the use of two stages of electrolysis. At the first stage of electrolysis, the anode product‑1 (17.3–48.5 % of the initial Pb-Bi alloy) of the composition has been isolated,%: 16.6–48.4 Bi; 51.4–83.2 Pb; operational extraction,%: 92.2–96.6 Bi; 9.8–44.4 Pb; main phases Bi0,3Pb0,7 and Bi0,95Pb0,05. A six-fold bismuth enrichment is achieved in the anode product. At the second stage of electrolysis of the previously isolated anode product of the composition,%: 26.7 Bi; 73.1 Pb; 0.13 Cu; 0.08 Zn, the anode product‑2 (28.1 % of the enriched Pb-Bi alloy) of the composition has been separated,%: 93.6 Bi; 4.1 Pb; 0.086 Ag; 0.0066 As; 0.006 Sb; 0.0013 Cu; 0.001 Sn; 0.0014 Zn; stage extraction,%: 98.6 Bi; 1.6 Pb; main phase Bi0,95Pb0,05. As a result of pyroelectrometallurgical processing of a Pb-Bi alloy (~10 % Bi) with anode polarization in two stages, an anode product (8.7 % of the initial alloy) of the composition has been isolated,%: ≥ 93.6 Bi; 4.1 Pb; extraction from the initial alloy,%: 93.0 Bi; 0.4 Pb has been obtained. The following modes are recommended for pyroelectrometallurgical processing in two stages of Pb-Bi alloy: process temperature 550–600 °C; anode current density: 0.5 A/cm2 at the first stage; 0.2–0.3 A/cm2 at the second stage; cathode current density: 1.5 A/cm2 at the first stage; 1.0 A/cm2 at the second stage; operating voltage on the tub: at the first stage 8–12 V; at the second stage 5–8 V; the composition of the electrolyte at both stages,%: 7 NaCl; 35 KCl; 18 PbCl2; 40 ZnCl2; the amount of electrolyte output for processing: at the first stage – 10 % of the mass of the Pb-Bi alloy after alkaline treatment; at the second stage – 10 % of the mass of the anode product of the first stage

Selective Recovery of Bismuth in Copper Electrolyte Through Coprecipitation Method and Its Mechanism

Selective Recovery of Bismuth in Copper Electrolyte Through Coprecipitation Method and Its Mechanism

Development of a Single-Step Microwave-Assisted Digestion Method using Dilute Nitric Acid for Determination of Bismuth in Bismuth-containing Pharmaceuticals by Hydride Generation-Atomic Fluorescence Spectrometry

A single-step microwave assisted digestion (MWAD) procedure employing very dilute solutions of HNO3 was developed for the quantitative determination of bismuth in bismuth-containing pharmaceuticals by hydride generation-atomic fluorescence spectrometry (HG-AFS). Experimental parameters affecting MWAD process such as acid concentration (HNO3), digestion time and temperature were optimized to get quantitative recovery of bismuth. The studies indicated that the method is rapid (within 15 min) including cooling time and recovery > 98% was obtained using 10mL of 5% (v/v) HNO3 as digestion medium with ~0.1g of sample. The optimum microwave digestion parameters obtained were: temperature – 180oC, pressure – 25 bar and hold time - 10 min. A clear solution with negligible residue was obtained after microwave digestion. The digested sample solution was appropriately diluted with 2% (v/v) HCl for subsequent analysis by HG-AFS. The reproducibility, expressed as % RSD was lower than 2% for the allopathic medicine. Under optimal conditions, the limit of detection (LOD) for Bi was calculated to be 0.024mg/kg. The methodology was optimized using a bismuth-containing pharmaceutical – Pylobis, purchased from a local pharmacy. The optimized MWAD approach was further applied to few other bismuth-containing pharmaceutical products. The developed method has significant advantages when compared to the conventional hot-plate digestion methods reported for Bi-containing pharmaceuticals, employing large volumes of concentrated acids. These investigations revealed that the proposed MWAD method in combination with HG-AFS can be utilized for the rapid determination of Bi in pharmaceutical products on regular basis.

Removal of Antimony and Bismuth from Copper Electrorefining Electrolyte: Part II—An Investigation of Two Proprietary Solvent Extraction Extractants

Antimony and bismuth recovery from copper electrorefining electrolyte could reduce the impacts of these problem elements and produce a new primary source for them. Two proprietary phosphonic acid ester extractants were examined (REX-1 and REX-2) for the removal of antimony and bismuth from copper electrorefining electrolytes. Experimentation included shakeout and break tests to determine the basic parameters for the extractants in terms of maximum loading, break times, and extraction and stripping efficiency. Five permutations of extractant mixtures (100 wt.% REX-1 and 25 wt.%, 50 wt.%, 75 wt.% and 100 wt.% REX-2) were studied. It was determined that REX-2 was able to extract Sb and Bi from the electrolyte, but required some mixture with REX-1 to better facilitate stripping with 400 g/L sulfuric acid. The laboratory electrorefining electrolyte containing glue had faster disengagement times than a synthetic solution without glue.

Effect of H2O2 on the Separation of Mo-Bi-Containing Ore by Flotation

Hydrogen peroxide (H2O2) is a strong oxidizer that causes non-selective oxidation of sulfide minerals, and its influence on bismuth sulfide ores is not well-documented. In this study, H2O2 was proposed as an alternative bismuthinite depressant, and its effect on a Mo-Bi-containing ore was intensively investigated by batch flotation tests. Results showed that the addition of H2O2 significantly destabilized the froth phase, thus decreasing the solids and water recovery. The recovery of bismuth in molybdenum concentrate was dramatically decreased to 4.64% by H2O2 compared with that in the absence of H2O2 (i.e., 50.14%). The modified first-order kinetic model demonstrated that the flotation rate of molybdenite slightly declined after H2O2 addition, whereas that of bismuthinite was drastically reduced from 0.30 min−1 to 0.08 min−1 under the same condition. Simulation revealed that H2O2 affected the floatability of both molybdenite and bismuthinite but resulted in more detrimental effect to bismuthinite. Hence, H2O2 has the potential to act as an effective depressant in bismuth sulfide ore flotation.

Determination of trace bismuth by atomic fluorescence spectrometry in blood

Objective: To establish a method for determination trace bismuth in blood by atomic fluorescence spectrometry (AFS) . Methods: 4.0 ml nitric acid and 1.0 ml perchloric acid was added into 1.0 ml blood sample then through automation graphite digestion instrument digested, after that 1.0 ml thiocarbamide-vitamin (10%) was injected, 8% HCl constant volume to 10.0 ml, the bismuth was detected by atomic fluorescence spectrometry with 5.0 ml digestive sample. Results: The method showed a linear relationship within the range of 0.4-50.0 μg/L (r=0.999 7) . The within-run and between-run relative standard deviations (RSD) of repetitive measurement at 10.0, 20.0, 40.0 μg/L concentration levels were 2.2%-4.9% and 3.0%-4.0%. The detection limit was 0.032 μg/L. The recoveries of bismuth were 93.0%-103.9%. Conclusion: This method is low detection limit, good accurate and high sensitivity. It has been applied for determination of trace bismuth in blood samples those who need occupation health examination or poisoning diagnosis.

Determination of bismuth in different samples by dispersive liquid–liquid microextraction coupled with microvolume β-correction spectrophotometry

A novel microvolume β-correction spectrophotometric method was proposed for the sensitive and selective determination of bismuth coupling with dispersive liquid–liquid microextraction (DLLME). The basis of the method is a quantitative colorimetric reaction between Bi3+ ions and dithizone, and the orange–brown colored complex of Bi(HDz)3 was then extracted into CCl4 phase by DLLME. Parameters related to the efficiency of microextraction, such as pH, complexant concentration, the volume ratio of disperser solvent and extraction solvent were discussed and optimized in detail. Under the optimized conditions, the real absorbance was in proportion to bismuth concentration in the range of 3–100μg·L−1 with a correlation coefficient (R) of 0.993. The limit of detection (LOD) and limit of quantitation (LOQ) were 0.54μg·L−1 and 1.80μg·L−1, respectively. Good recoveries of bismuth were obtained in the range of 96.3–103.6% in different real samples confirming the accuracy of developed procedure and its independence from matrix interference.

Processes for the recovery of bismuth from ores and concentrates

Some data on the bismuth recovery technologies in the production of copper and the processing of copper-bismuth and tungsten-molybdenum concentrates are given.

Separation and Extraction of Bismuth and Manganese from Roasted Low-Grade Bismuthinite and Pyrolusite: Thermodynamic Analysis and Sulfur Fixing

A new environmentally friendly technology with higher recovery of bismuth is proposed to extract bismuth from low-grade bismuthinite and co-production MnSO4 from low-grade pyrolusite. The effects of simultaneous roasting process parameters on the sulfur-fixing rate and MnSO4 formation rate are investigated. Based on the Pourbaix diagram of metal-sulfur-oxygen system, the behavior of bismuth, manganese, and associated metal elements such ferrous, copper, lead, and sulfur in the bismuthinite and pyrolusite during roasting process is analyzed. The experimental results show that Bi in the ores can be converted into bismuth oxide or oxygen bismuth sulfate, and most of Mn in the ores can react with SO2 from bismuthinite to form MnSO4, which agree with thermodynamic analysis. A maximum of sulfur-fixing rate of 98.14% and MnSO4 formation rate of 70.2% are obtained under the conditions of 1.4 for the molar ratio of MnO2 to total sulfur in mixing ores of bismuthinite and pyrolusite (n(MnO2)/n(S)), 923 K for the roasting temperature, 2 h for roasting time, 140 L/h for air rate, and less than 74 μm for particle size. The ultimate recovery rate of bismuth reaches 96.25% by selective leaching of the roasted product, purification of leaching solution, and hydrolysis, which is higher than the current applied technology for the low-grade bismuthinite.

Selective leaching and recovery of bismuth as Bi2O3 from copper smelter converter dust

Secondary materials such as converter dusts from the copper smelters are rich in bismuth and other valuable metals. This paper describes an economically viable Bi recovery process from the converter dust by a hydrometallurgical method. The results showed that, Bi could be effectively leached out from the converter dust using a combination of H2SO4 and NaCl as leaching reagent with a leaching efficiency of 92% under the optimum conditions. Small amount of impurities such as Ag, Pb, Te and Se were removed from the leach solution by SO2 reduction. Bismuth was separated from the leach solution as crude BiOCl cake by hydrolysis method. The final bismuth product, Bi2O3 was synthesized by NaOH treatment of the crude BiOCl cake at elevated temperature. The purity of the Bi2O3 product was about 97.8%, As, Pb and Sb being the major impurities. Based on the feasibility study, the process was deemed technically feasible and its product commercially saleable.

Electrolytic recovery of bismuth and copper as a powder from acidic sulfate effluents using an emew® cell

Effective removal of bismuth is a primary concern during copper electrorefining.

Direct Determination of Bismuth in Solid Samples by HG-FAAS

Current work deals with direct determination of Bismuth element insolid samples using a labmade Hydride Generation(HG) accessory for Flame Atomic Absorption Spectrometry (FAAS) instrument. In this method an appropriate amount of solid powder sample, are added into the HG reactor and the produced bismuth trihydride is then transferred into the slotted quartz tube just above the flame where analysis is performed. The samples of viscose and liquid solutions can be analyzed after drying the sample in HG reactor prior to generating the hydride vapor. Affecting parameters including: carrier gas flow rate, volume and concentration of the HCl acid and reducing agent solution were optimized to achieve best results. Dynamic range of calibration was linear between 20 to 2500 ppb and Relative Standard Deviation (RSD) for 4 times measurement was 5.44%. The recovery of Bismuth in real samples of Bismuthsubcitrate tablets and Bismuthsubnitrate ointment by this method were found to be 98.14% and 84.08%,

Solubility of Bismuth in Liquid Bi-S Based Free Cutting Steel

AbstractSolubility of bismuth in liquid Bi-S based free cutting steel was measured using a vapor-liquid equilibration method at 1540–1600 °C, and the recovery rate of bismuth in the steel with different temperatures under an atmospheric pressure was also measured. The results showed that the solubility of bismuth in liquid Bi-S based free cutting steel from experiment under a constant volume at 1540, 1560, 1580, and 1600 °C were 0.174, 0.181, 0.205, and 0.220 mass%, respectively, and the relationship of bismuth solubility vs. temperature could be expressed as lg[%Bi] = −6049/T + 2.572. Meanwhile, the solubility of bismuth increased with the increase of Mn content, but decreased with the increase of C content. The recovery of bismuth in this experiment reached a maximum when the temperature was at bismuth boiling point or so, and then it was decreased with the increase of temperature when the temperature was above 1560 °C, which might be attributed to the accelerating of bismuth evaporation that were caused by the increase of bismuth equilibrium partial pressure above the surface of the molten steel with increasing temperature.