A method for recovering antimony alkali residue

Abstract

This application provides a method for recovering antimony alkali residue. This process combines microwave activation with citric acid leaching technology to achieve highly efficient antimony recovery (>95%) while reducing energy consumption by 40%. Through TBP gradient extraction, a 98% recovery rate is achieved, eliminating the generation of hazardous waste. Zero waste is achieved, with tailings completely converted into qualified building materials and a wastewater reuse rate exceeding 95%. The process offers significant economic benefits, with a treatment cost of approximately 1,200 yuan per ton and an investment payback period of less than three years. The process utilizes mature industrial equipment, enabling large-scale application. A closed-loop design ensures process stability. Comprehensive safety measures are implemented, ensuring that the final product meets national environmental protection and building material standards.

Description

Technical Field

[0001] The present invention relates to the field of metallurgy technology, and in particular to a method for recovering antimony alkali slag. Background Art

[0002] Antimony alkali slag is a typical hazardous solid waste generated during antimony smelting, lead and zinc smelting, and electronic waste processing. Its main components are antimony oxide (Sb2O3 / Sb2O5), silicates, heavy metals (arsenic, lead, cadmium), and residual alkali. With the growth of global antimony consumption (an average annual growth rate of approximately 4.2%), antimony alkali slag stockpiles continue to climb, with annual production in China alone exceeding 500,000 tons. Existing antimony alkali slag treatment technologies generally suffer from low recovery efficiency, high risk of secondary pollution, and a limited resource recovery pathway, severely hindering the sustainable development of the industry. Traditional processes mostly use high-temperature roasting (600-800°C) to break the silicate encapsulation structure, but this process relies on external heat conduction and has significant drawbacks: excessive energy consumption, with roasting one ton of slag consuming 1200-1500 kWh of electricity, accounting for more than 60% of the total energy consumption of the entire process, and high temperatures causing partial volatilization of antimony oxides (loss rate of 5-8%); low activation efficiency, with silicate minerals only dissociating on the surface under static heating, making it difficult to release internal antimony, and subsequent leaching rates limited to below 80%; pollution spreads, with high-temperature environments exacerbating the oxidation and volatilization of arsenic (the boiling point of As2O3 is 460°C), generating highly toxic arsenic-containing smoke, requiring the configuration of a complex exhaust gas purification system, and increasing pollution control costs (approximately ￥500 / ton of slag).

[0003] Mainstream hydrometallurgical technology uses a full leaching strategy with strong acid (6-8 mol / L HCl or H2SO4). Although this can extract antimony, it faces multiple bottlenecks: poor selectivity. High acidity leads to the simultaneous dissolution of heavy metals such as arsenic, lead, and cadmium (As leaching rate >40%), resulting in a surge in subsequent purification loads and the need for multi-stage sulfide precipitation (Na2S or FeS). This generates hazardous arsenic-containing waste (As2S3 content >15%), with disposal costs as high as ¥800-1000 per ton. Acid mist is a hazard. Hydrochloric acid volatilization produces Cl2 and HCl gases, creating a harsh operating environment (requiring corrosion-resistant and sealed equipment). Acid consumption reaches 0.8-1.2 tons per ton of slag, and reagent costs account for over 35%. Antimony is lost through hydrolysis. Traditional processes ignore dynamic pH control. When the leachate pH exceeds 1.5, antimony easily hydrolyzes to form SbOCl precipitate (with a solubility of only 0.02 g / L), resulting in a 10-15% decrease in antimony recovery. Existing separation technologies, primarily sulfide precipitation-pyrometallurgical refining, suffer from significant shortcomings: multi-stage precipitation is inefficient. To remove impurities such as As and Pb, sodium sulfide and lime milk must be added sequentially to adjust the pH, a process that takes >8 hours. Antimony and impurities experience significant co-precipitation (loss rate of 12-18%). The product's added value is low, with the final product being antimony concentrate (70-85% Sb), which requires secondary smelting (1200°C) to purify to above 99%, increasing energy consumption by an additional 800-1000 kWh / ton. Hazardous waste accumulates in a chain reaction, generating 300-400 kg of arsenic-containing filter residue, 8-10 m³ of acidic wastewater, and 50-80 kg of waste gas purification residue per ton of slag. The overall hazardous waste rate exceeds 45%, far exceeding the limit specified in the National List of Hazardous Wastes (category HW48).

[0004] Traditional processes for treating leached residues are crude and pose significant environmental risks. Residual heavy metals (As 5-8%, Pb 0.5-1.2%) in landfill residues are susceptible to leakage due to rainwater leaching, with leachate As concentrations reaching 2-5 mg / L (20-50 times the limit set in GB 5085.3-2007). Some companies use cement for simple solidification, but the silicate matrix has a weak ability to fix As³⁺ / Sb³⁺ (solidification rate <70%) and poor long-term stability. Wastewater treatment costs are high, as acidic wastewater requires a multi-stage treatment process involving neutralization, precipitation, and membrane filtration, costing ¥30-50 per ton of water. Furthermore, membrane fouling and clogging are frequent, and the reuse rate is less than 50%. Existing technologies are difficult for small and medium-sized enterprises to bear due to complex processes and large equipment investments (rotary kiln + multi-stage purification system > ￥50 million / 10,000 ton line): costs are inverted, with the comprehensive cost of processing a ton of slag being ￥1,800-2,200, while the selling price of antimony concentrate is only ￥15,000-20,000 / ton (the market price of metallic antimony is ￥30,000 / ton), with a profit margin of <10%; policy risks, environmental protection taxes (￥1,000-1,500 / ton hazardous waste) and carbon emission costs (about ￥200 / ton CO2) further compress profit margins; technology is locked in, and the iteration of traditional processes has stagnated. In the past decade, 90% of patents have been local optimizations (such as leaching agent replacement), lacking systematic innovation.

[0005] Existing antimony alkali slag treatment technologies are limited by core flaws such as high energy consumption, low selectivity, the generation of hazardous waste chains, and poor economic efficiency, making them unable to meet the green smelting needs under the "dual carbon" goals. The industry urgently needs to develop a new, efficient, clean, fully resource-recoverable, and low-cost recycling process to break the "recycling-equals-pollution" dilemma and promote the antimony industry's transition to a circular economy model. Summary of the Invention

[0006] In order to solve or partially solve the problems existing in the related art, the present invention provides a method for recovering antimony alkali slag.

[0007] The following steps are involved:

[0008] (1) Microwave activation pretreatment: crush the antimony alkali slag to a particle size of ≤0.2 mm, and calcine it at 300-400℃ for 10-20 minutes under the conditions of microwave frequency of 2.0-3.0 GHz and power of 3-7 kW;

[0009] (2) Acidic oxidation leaching: the product of step (1) is mixed with a mixed acid solution at a liquid-to-solid ratio of 5:1-7:1, wherein the mixed acid is hydrochloric acid and citric acid at a volume ratio of 3:1-5:1, 0.3%-0.7% H2O2 is added, and leaching is carried out at 75-85°C and pH 1.4-1.6 for 1.5-2.5 hours;

[0010] (3) Gradient extraction separation: adjust the pH of the leachate to 1.8-2.2, add TBP-kerosene extractant according to the ratio of organic phase to aqueous phase = 1:3-1:5, and use 3%-7% sulfuric acid solution for back extraction;

[0011] (4) Antimony extraction by electrodeposition: electrolyze the stripping solution at a current density of 150-250 A / m², a cell voltage of 2.5-3.5 V, and control the temperature at 30-50°C;

[0012] (5) Tailings resource utilization: leaching residues are mixed with auxiliary materials at a mass ratio of 0.5:1-1.5:1, 5%-15% binder is added, and after pressing and molding, the mixture is cured in an environment with a humidity of ≥85%;

[0013] (6) Waste liquid recycling: The waste acid is vacuum distilled to recover HCl with a concentration of ≥80%, and the waste water is neutralized and filtered through a membrane for reuse.

[0014] Furthermore, in step (1), the microwave power is 5 kW, the roasting temperature is 380° C., and the roasting time is 15 minutes.

[0015] Furthermore, in step (2), the liquid-to-solid ratio is 6:1, the volume ratio of hydrochloric acid to citric acid is 4:1, the amount of H2O2 added is 0.5%, the leaching temperature is 80°C, and the pH is 1.5.

[0016] Furthermore, in step (3), the volume ratio of the organic phase to the aqueous phase is 1:4, and the concentration of the stripping sulfuric acid is 5%.

[0017] Furthermore, in step (4), the current density is 200 A / m², the cell voltage is 3.0 V, and the electrolysis temperature is 40°C.

[0018] Furthermore, the auxiliary material in step (5) is fly ash, the mass ratio of residue to fly ash is 1:1, the binder is cement and the addition amount is 10%, the molding pressure is 15 MPa, and the curing humidity is >90%.

[0019] Furthermore, the purity of the recovered HCl in step (6) is ≥85%, and the membrane filtration adopts a nanofiltration membrane with a molecular weight cut-off of 200-400 Da.

[0020] Furthermore, in step (3), the volume ratio of TBP to kerosene is 1:4, and the stripping time is 15 minutes.

[0021] Furthermore, in step (5), sodium sulfide is used for precipitation of heavy metal sulfides, and the addition amount is 0.2 kg / kg slag.

[0022] Furthermore, the antimony ion concentration of the leachate in step (2) is ≥50 g / L, and the cathode antimony purity in step (4) is ≥99.5%.

[0023] It should be understood that the foregoing general description and the following detailed description are exemplary and explanatory only and are not restrictive of the present invention.

[0024] Beneficial technical effects of the present invention:

[0025] This process combines microwave activation and citric acid leaching technology to achieve efficient antimony recovery (>95%) while reducing energy consumption by 40%; a 98% recovery rate is achieved through TBP gradient extraction, avoiding the generation of hazardous waste; zero waste emissions are achieved, tailings are completely converted into qualified building materials, and the wastewater reuse rate is over 95%; the economic benefits are significant, with a treatment cost of approximately 1,200 yuan / ton and an investment payback period of less than 3 years; mature industrial equipment is used and conditions for large-scale application are met; a closed-loop circulation design ensures process stability; and complete safety measures are equipped, so that the final product meets national environmental protection and building materials standards. DETAILED DESCRIPTION

[0026] The following describes alternative embodiments of the present invention in more detail. Although alternative embodiments of the present invention are described, it should be understood that the present invention can be implemented in various forms and should not be limited by the embodiments described herein. Instead, these embodiments are provided to make the present invention more thorough and complete, and to fully convey the scope of the present invention to those skilled in the art.

[0027] The present invention provides a method for recovering antimony alkali residue.

[0028] The following steps are involved:

[0029] (1) Microwave activation pretreatment: crush the antimony alkali slag to a particle size of ≤0.2 mm, and calcine it at 300-400℃ for 10-20 minutes under the conditions of microwave frequency of 2.0-3.0 GHz and power of 3-7 kW;

[0030] Microwaves excite polar molecules (such as H₂O and SiO⁻) in the antimony alkali slag through a high-frequency electromagnetic field (2.45 GHz), triggering violent molecular vibration and friction, rapidly generating heat energy (300–400°C). This endogenous heating effectively destroys the silicate encapsulation structure, exposing and activating the antimony oxides (Sb₂O₃ / Sb₂O₅), while also avoiding the energy waste caused by uneven heat conduction during traditional roasting.

[0031] (2) Acidic oxidation leaching: the product of step (1) is mixed with a mixed acid solution at a liquid-to-solid ratio of 5:1-7:1, wherein the mixed acid is hydrochloric acid and citric acid at a volume ratio of 3:1-5:1, 0.3%-0.7% H2O2 is added, and leaching is carried out at 75-85°C and pH 1.4-1.6 for 1.5-2.5 hours;

[0032] In mixed acid (HCl + citric acid), H⁺ dissolves antimony oxide to form Sb³⁺ / Sb 5 ⁺ ions, while citric acid complexes antimony ions through carboxyl groups, inhibiting their hydrolysis and precipitation, while reducing HCl volatilization. H2O2 acts as an oxidant to oxidize part of Sb³⁺ to Sb 5 ⁺ (reaction formula: Sb2O3 + H2O2 → Sb2O5 + H2O), improving the leaching rate of antimony. By precisely controlling the pH to 1.5, the dissolution of arsenic (As³⁺ is readily soluble at pH>2) and lead (Pb²⁺ precipitates as PbCl2 at low pH) is suppressed, achieving selective separation of antimony.

[0033] (3) Gradient extraction separation: adjust the pH of the leachate to 1.8-2.2, add TBP-kerosene extractant according to the ratio of organic phase to aqueous phase = 1:3-1:5, and use 3%-7% sulfuric acid solution for back extraction;

[0034] Tributyl phosphate (TBP) reacts with Sb in acidic medium via phosphorus oxygen. 5 ⁺ forms a hydrophobic complex (reaction: SbCl⁻ + 3TBP → [SbCl⁻·3TBP]), which is preferentially extracted into the organic phase, while impurities such as Cu²⁺ and Zn²⁺ remain in the aqueous phase due to their weak complexing ability. During stripping, dilute sulfuric acid (H₂SO⁄) provides a high concentration of H⁺, which breaks down the complex structure ([SbCl⁃·3TBP] + H₂SO⁄ → Sb³⁺ + 3TBP + SO⁻²⁻ + Cl⁻), releasing antimony ions into the aqueous phase, and the organic phase is regenerated and recycled.

[0035] (4) Antimony extraction by electrodeposition: electrolyze the stripping solution at a current density of 150-250 A / m², a cell voltage of 2.5-3.5 V, and control the temperature at 30-50°C;

[0036] Under the influence of a DC electric field, Sb³⁺ in the stripping solution is reduced to metallic antimony on the cathode (titanium plate) surface (reaction: Sb³⁺ + 3e⁻ → Sb), while the lead anode undergoes an oxidation reaction (Pb → Pb²⁺ + 2e⁻). By controlling the current density (200 A / m²) and temperature (40°C), the antimony crystal morphology is optimized, short-circuiting caused by dendrite growth is avoided, and the cathode antimony purity is guaranteed to exceed 99.5%.

[0037] (5) Tailings resource utilization: leaching residues are mixed with auxiliary materials at a mass ratio of 0.5:1-1.5:1, 5%-15% binder is added, and after pressing and molding, the mixture is cured in an environment with a humidity of ≥85%;

[0038] Na2S is added to the residual leachate to form insoluble sulfides (such as CuS, ZnS, solubility product Ksp≈10⁻³) with Cu²⁺ and Zn²⁺. 6), and is rendered harmless through sedimentation and separation.

[0039] The leached residue (SiO2, CaO) and the active Al2O3 in the fly ash form a dense structure in the cement hydration reaction (Ca(OH)2 + SiO2 →CSH gel), which physically encapsulates and chemically solidifies the residual heavy metals.

[0040] (6) Waste liquid recycling: The waste acid is vacuum distilled to recover HCl with a concentration of ≥80%, and the waste water is neutralized and filtered through a membrane for reuse.

[0041] Vacuum distillation (80°C) recovers HCl gas through the boiling point difference between HCl and water (HCl boiling point -85°C, water 100°C), and then condenses and reuses it, with an acid recovery rate of >85%.

[0042] In one embodiment of the present application, in step (1), the microwave power is 5 kW, the roasting temperature is 380° C., and the roasting time is 15 minutes.

[0043] In one embodiment of the present application, in step (2), the liquid-to-solid ratio is 6:1, the volume ratio of hydrochloric acid to citric acid is 4:1, the amount of H2O2 added is 0.5%, the leaching temperature is 80°C, and the pH is 1.5.

[0044] In one embodiment of the present application, in step (3), the volume ratio of the organic phase to the aqueous phase is 1:4, and the concentration of the stripping sulfuric acid is 5%.

[0045] In one embodiment of the present application, in step (4), the current density is 200 A / m², the cell voltage is 3.0 V, and the electrolysis temperature is 40°C.

[0046] In one embodiment of the present application, the auxiliary material in step (5) is fly ash, the mass ratio of residue to fly ash is 1:1, the binder is cement and the addition amount is 10%, the molding pressure is 15 MPa, and the curing humidity is >90%.

[0047] In one embodiment of the present application, the purity of the HCl recovered in step (6) is ≥85%, and the membrane filtration adopts a nanofiltration membrane with a molecular weight cut-off of 200-400 Da.

[0048] In one embodiment of the present application, in step (3), the volume ratio of TBP to kerosene is 1:4, and the stripping time is 15 minutes.

[0049] In one embodiment of the present application, sodium sulfide is used for precipitation of heavy metal sulfides in step (5), and the addition amount is 0.2 kg / kg slag.

[0050] In one embodiment of the present application, the antimony ion concentration of the leachate in step (2) is ≥50 g / L, and the cathode antimony purity in step (4) is ≥99.5%.

[0051] For the purpose of greater clarity, the present invention is described in detail with reference to the following examples.

[0052] Example 1: Microwave activation: crushed slag particle size 0.2 mm, microwave frequency 3.0 GHz, power 7 kW, temperature 400°C, calcination for 20 minutes;

[0053] Acid leaching: liquid-solid ratio 7:1, hydrochloric acid: citric acid = 5:1, H2O2 addition amount 0.7%, temperature 85℃, pH = 1.6, leaching 2.5 hours;

[0054] Gradient extraction: pH = 2.2, organic phase: aqueous phase = 1:3, back extraction sulfuric acid concentration 7%;

[0055] Electrodeposition: current density 250 A / m², cell voltage 3.5 V, temperature 50°C;

[0056] Tailings resource utilization: residue: fly ash = 1.5:1, cement addition 15%, pressure 17 MPa;

[0057] Waste liquid circulation: HCl recovery concentration is 80%, and the nanofiltration membrane has a molecular weight cutoff of 400 Da.

[0058] Example 2: Microwave activation: crushed slag particle size 0.15 mm, microwave frequency 2.45 GHz, power 5 kW, temperature 380°C, calcination for 15 minutes;

[0059] Acid leaching: liquid-solid ratio 6:1, hydrochloric acid: citric acid = 4:1, H2O2 addition amount 0.5%, temperature 80℃, pH = 1.5, leaching for 2 hours;

[0060] Gradient extraction: pH = 2.0, organic phase: aqueous phase = 1:4, back extraction sulfuric acid concentration 5%;

[0061] Electrodeposition: current density 200 A / m², cell voltage 3.0 V, temperature 40°C;

[0062] Tailings resource utilization: residue: fly ash = 1:1, cement addition 10%, pressure 15 MPa;

[0063] Waste liquid circulation: HCl recovery concentration is 85%, and the nanofiltration membrane has a molecular weight cutoff of 300 Da.

[0064] Example 3: Microwave activation: crushed slag particle size 0.1 mm, microwave frequency 2.0 GHz, power 3 kW, temperature 300°C, calcination for 10 minutes;

[0065] Acid leaching: liquid-solid ratio 5:1, hydrochloric acid: citric acid = 3:1, H2O2 addition amount 0.3%, temperature 75℃, pH = 1.4, leaching 1.5 hours;

[0066] Gradient extraction: pH = 1.8, organic phase: aqueous phase = 1:5, back extraction sulfuric acid concentration 3%;

[0067] Electrodeposition: current density 150 A / m², cell voltage 2.5 V, temperature 30°C;

[0068] Tailings resource utilization: residue: fly ash = 0.5:1, cement addition 5%, pressure 13 MPa;

[0069] Waste liquid circulation: HCl recovery concentration is 83%, and the nanofiltration membrane has a molecular weight cutoff of 200 Da.

[0070] Comparative analysis of results

[0071]

[0072] Test example

[0073] Raw materials: the same batch of antimony alkali slag (composition: Sb 12.5%, As 3.2%, Pb 1.8%, SiO2 35%, CaO 20%), processing capacity 100 kg;

[0074] Traditional crafts:

[0075] Calcination: rotary kiln 650℃×2 hours;

[0076] Leaching: 6 mol / L HCl, liquid-to-solid ratio 5:1, 80°C × 4 hours;

[0077] Precipitation: Na2S sulfide precipitation, the filter residue is dried to obtain antimony concentrate (Sb 85%);

[0078] Waste residue: hazardous waste landfill, neutralization discharge into wastewater.

[0079] Example 2 process:

[0080] Microwave activation (380℃×15 minutes) → selective leaching (HCl+citric acid) → gradient extraction (TBP) → electrodeposition → tailings brick making → closed-loop circulation of waste liquid.

[0081] Comparison of key indicators

[0082]

[0083] Economic benefit verification

[0084]

[0085] Test Example 2

[0086] Test Method

[0087] Sample: building bricks produced by the process of Example 2 (cured for 28 days);

[0088] Standard: Solid Waste Leaching Toxicity Leaching Method - Acetate Buffer Solution Method (HJ / T 300-2007);

[0089] Conditions: liquid-to-solid ratio 20:1, pH = 2.88 ± 0.05, tumbling and shaking for 18 hours;

[0090] Detection: ICP-MS was used to determine the concentration of the extract.

[0091] 2. Test Results

[0092]

[0093] Note: The measured values ​​are far lower than the national limit values ​​(As<1%, Pb<2.5%), and lower than the industrial application safety threshold of the groundwater Class III standard (As≤0.05mg / L, Pb≤0.05 mg / L).

Claims

1. A method for recovering antimony alkali residue, characterized in that: The following steps are involved: (1) Microwave activation pretreatment: The antimony alkali slag was crushed to a particle size of 0.15 mm and calcined at 380 °C for 15 min under the conditions of microwave frequency of 2.45 GHz and power of 5 kW; (2) Acidic oxidation leaching: the product of step (1) was mixed with a mixed acid solution at a liquid-to-solid ratio of 6:1, wherein the mixed acid was hydrochloric acid and citric acid at a volume ratio of 4:1, 0.5% H2O2 was added, and leaching was carried out at 80°C and pH 1.5 for 2 hours; (3) Gradient extraction separation: adjust the pH of the leaching solution to 2.0, add TBP-kerosene extractant at a ratio of organic phase to aqueous phase = 1:4, and use 5% sulfuric acid solution for back extraction; (4) Antimony extraction by electrodeposition: The stripping solution was electrolyzed at a current density of 200 A / m², a cell voltage of 3.0 V, and a temperature of 40°C. (5) Tailings resource utilization: the leaching residue is mixed with auxiliary materials in a mass ratio of 1:1, 10% binder is added, pressed and molded, and then cured in an environment with a humidity of >90%. The auxiliary material is fly ash, and the molding pressure is 15 MPa; (6) Waste liquid recycling: The waste acid is vacuum distilled to recover HCl with a concentration of ≥85%, and the wastewater is neutralized and then filtered for reuse. The membrane filtration uses a nanofiltration membrane with a molecular weight cutoff of 300 Da.

2. The method according to claim 1, wherein: In step (3), the volume ratio of TBP to kerosene is 1:4, and the stripping time is 15 minutes.

3. The method according to claim 1, wherein: The antimony ion concentration of the leachate in step (2) is ≥50 g / L, and the cathode antimony purity in step (4) is ≥99.5%.

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