Method for graded recovery of sulfur and metal in multi-metal sulfur slag by combination of wet process and pyrogenic process

Abstract

The invention discloses a method for graded recovery of sulfur and metal in multi-metal sulfur slag by combination of a wet process and a pyrogenic process, mainly relates to a method for realizing separation and recovery of valuable metal by regulating sulfur phase reconstruction by the wet process and cooperating with phase conversion of multi-metal of zinc, lead and silver by the pyrogenic process, and the method is convenient to operate, good in treatment effect and high in recovery rate. The method effectively realizes separation of metal and sulfur, realizes controllable conversion and refining of valuable metal phases of mercury, zinc, lead and silver, realizes conversion of sulfur elements into elemental sulfur, realizes separation of elemental sulfur, comprehensively recovers valuable metals and sulfur resources in hazardous wastes, and has significant economic benefits.

Description

Technical Field

[0001] The present invention relates to the field of nonferrous metal smelting, and in particular to a method for recovering sulfur and metals from polymetallic sulfur slag by combining a wet method with a pyrolytic method. Background Art

[0002] Polymetallic sulfur slag is usually a by-product produced during the smelting process, containing a variety of metals (such as mercury, zinc, lead, silver, etc.) and a large amount of sulfides. The recovery of sulfur is not only economically valuable, but also meets environmental protection requirements and can avoid the pollution of the environment by sulfides. The recovery of metals in sulfur slag is a greater challenge, especially the metal components in high-sulfur slag easily form sulfides, and these sulfides are difficult to recover separately by conventional wet or pyrometallurgical methods. Therefore, how to recycle these valuable resources efficiently and environmentally friendly has become a difficult problem in the field of non-ferrous metal smelting. Existing wet recovery mainly includes pressurized oxygen leaching, which mainly uses oxygen to oxidize sulfides under high temperature and high pressure conditions to achieve the separation of sulfur and metals. Specifically, by adjusting the oxygen partial pressure, temperature, and acidity to suppress excessive oxidation, sulfur is recovered as a single substance (S 0) form, and the metal is mainly recovered through subsequent extraction and electrolysis. However, it often has many defects, such as low sulfur purity (arsenic / organic matter content, <90%), severe equipment corrosion, and high risk of high-pressure operation. In addition, the bioleaching method mainly uses the oxidizing ability of specific microorganisms to dissolve the metals in the sulfur slag, and then recovers the metals through extraction and electrolysis. However, it has problems such as long treatment cycle (>30 days), temperature sensitivity (25-45℃), and weak bacterial resistance to toxicity (inactivation when As>5g / L), which also limit its large-scale application. Traditional pyrometallurgical recovery technology primarily involves roasting and smelting to recover sulfur and metals. This involves roasting sulfur slag at 600-800°C, generating flue gases for acid production. The resulting roasted sand is then smelted at 1300°C to produce matte (Cu / Ni enriched), which also produces waste slag. Consequently, this method inevitably suffers from low sulfur recovery (10-20% sulfur escapes during roasting), difficulty capturing volatile metals (Pb / Zn / As) in the flue gases, high energy consumption (>1200 kWh / t of slag), SO2 emission risks, and difficult waste slag disposal. Furthermore, traditional wet processes focus on a "room temperature-solution reaction," using leaching agents to dissolve metal ions. Sulfur is typically discarded as an "inert residue" or requires additional treatment (such as high-temperature roasting for desulfurization). Technicians often assume that "after the wet process, sulfur has become a stable slag phase, and the pyrometallurgical process only requires metal treatment," overlooking the possibility that sulfur can be pre-activated during the wet process. The pyrometallurgical method emphasizes "high temperature-oxidation / reduction", and the removal of sulfur relies on combustion to form SO2 (which requires supporting desulfurization). Technicians prefer to improve the metal recovery rate by optimizing the roasting atmosphere (such as oxygen-rich / oxygen-poor) rather than considering "wet method to convert sulfur into non-gaseous products in advance" to avoid SO2 generation. Therefore, the combination of pyrometallurgical and wet methods is considered to convert the difficult-to-separate elemental sulfur from the "metal sulfide-encapsulated state" to the "free state / soluble state", breaking the dilemma of the traditional wet method of "metal-sulfur co-leaching". At the same time, the "clean sulfur slag" after treatment no longer contains elemental sulfur, but only metal sulfides. Only a small amount of SO2 is generated at high temperature, reducing the cost of tail gas treatment. In addition, the "free state / soluble state" can be reconstructed and recycled into high-purity S (purity>99%), turning waste into treasure. In view of this, the wet method and the pyrolysis method are combined to construct a graded recovery technology, combining the respective advantages of the two methods and giving full play to the advantages of the two methods. It is possible to efficiently extract valuable metals and sulfur from polymetallic sulfur slag in a relatively short time, which can not only improve the resource recovery rate but also reduce pollution to the environment. Summary of the Invention

[0003] The purpose of the present invention is to overcome the problem mentioned in the above background technology that it is difficult to achieve efficient separation of metal phase and elemental sulfur in multi-metal sulfur slag through a single wet method and fire method, and provide a method for separating and recovering valuable metals by regulating sulfur phase reconstruction by wet method and combining fire method with zinc, lead and silver multi-metal phase transformation. The method is easy to operate and has good processing effect, effectively realizes the separation of metal and sulfur, and simultaneously realizes the controllable transformation and refining of mercury, zinc, lead and silver valuable metal phases, converts sulfur element into elemental sulfur for separation, and comprehensively recovers valuable metals and sulfur resources in hazardous waste, with significant economic benefits. To achieve the above purpose, the present invention is constructed according to the following technical solutions:

[0004] A method for recovering sulfur and metals from polymetallic sulfur slag by combining wet and pyrolysis, comprising the following steps:

[0005] S1, mixing the polymetallic sulfur slag with a chelating leaching agent, performing a leaching reaction, and obtaining a sulfur leachate and a polymetallic residue after solid-liquid separation, wherein the sulfur leachate is reconstituted into sulfur by adjusting the pH;

[0006] S2, drying and crushing the multi-metallic residue, calcining to obtain a mercury-removed separation slag, and cooling and collecting the tail gas during the calcination process to obtain elemental mercury;

[0007] S3, crushing and calcining the demercured separation slag to obtain demercured metal-enriched slag;

[0008] S4, drying and crushing the demercured metal-enriched slag, performing a leaching reaction with an acid solution, obtaining a zinc leachate and dezincified metal-enriched slag after solid-liquid separation, and then neutralizing and replacing the zinc leachate, purifying and electrolyzing it to obtain elemental zinc;

[0009] S5, drying and crushing the dezincified metal-enriched slag, and then calcining it to obtain silver-containing crude lead, which is then refined by electrolytic refining, wherein the anode mud is acid-dissolved and then electrolyzed to obtain elemental silver.

[0010] The chelating leaching agent in S1 includes at least one of sodium sulfide, potassium sulfide, ammonium sulfide, sodium sulfate, and sodium hydroxide. The concentration of the chelating leaching agent is 10-20 wt%, the liquid-to-solid ratio is 10-20 L / kg, the leaching time is 0.5-4 hours, the temperature is 25-80° C., and the stirring speed is 100-800 rpm / min.

[0011] The formation of polysulfides is a "neutralization" reaction of the valence state of the S element. Sulfur salts of different valence states may form polysulfides with elemental sulfur, mainly due to the low-valence sulfur ion (S 2- / HS -) drives the nucleophilic attack on elemental sulfur, triggering the disproportionation reaction of elemental sulfur and the breakage and recombination of sulfur chains. This is a process involving redox (disproportionation) and chain growth (polymerization), ultimately forming polysulfide anions containing sulfur chains of different lengths [S n ] 2- / HS n -.

[0012] Furthermore,

[0013] In the step S1, the pH is adjusted to a range of 7.5 to 12.5, and the pH regulator is at least one of sulfuric acid and hydrochloric acid, with a concentration of 5 to 30 wt%.

[0014] In S2, the multi-metal residue is dried, crushed, passed through a 100-mesh sieve, and then calcined. The calcination atmosphere is at least one of air and oxygen, the temperature range is 300-600° C., and the calcination time is 1-5 hours.

[0015] In S3, the mercury removal separation slag is crushed and passed through a 100-mesh sieve before being calcined. The calcination atmosphere is at least one of air and oxygen. The calcination temperature ranges from 400 to 500 °C and the calcination time is 1 to 5 h.

[0016] The calcination in S3 is mainly to regulate the transformation of the metal phase in the slag, mainly by utilizing the fact that the metal sulfides in the slag will be transformed into the oxidized phase under high temperature and oxygen-rich conditions, thereby facilitating the subsequent acid leaching treatment.

[0017] In S4, the demercured metal-enriched slag is dried, crushed, passed through a 100-mesh sieve, and then used for acid leaching. The acid leaching temperature range is 60-120° C., the acid solution is at least one of sulfuric acid, hydrochloric acid, and nitric acid, the concentration is 10-30 wt%, and the leaching time is 1-5 hours.

[0018] In the step S4, the zinc leaching solution is adjusted to a pH of 1 to 4, so that iron and arsenic are separated and removed by forming insoluble compounds. Subsequently, the temperature is adjusted to 60 to 85°C and 1.5 to 2 times the theoretical amount of zinc powder is added to replace the copper, cadmium and other impurities. The electrolytic zinc smelting electrolytic cell structure is composed of an aluminum plate as a cathode and a crude lead containing 0.5 to 1% silver as an anode. The solution after zinc powder replacement and separation to remove copper, cadmium and other impurities is used as the electrolyte. The current density is 400 to 600 A / m 2 , the electrolysis voltage is 1~4V, and the electrolysis time is 3~10h.

[0019] In S5, the dezincified metal-enriched slag is dried and crushed, passed through a 100-mesh sieve, and then calcined. The calcination atmosphere is oxygen, and the temperature is 1100-1500° C. The electrolytic lead smelting cell structure is silver-containing crude lead as the anode, stainless steel as the cathode, 10-20wt% lead fluoride silicate solution as the electrolyte, and the current density is 100-200A / m 2 , the electrolysis voltage is 1~3V, and the electrolysis time is 2~5h.

[0020] The acid solution for dissolving the anode mud in S5 includes: at least one of sulfuric acid and nitric acid, with a concentration of 10-20 wt%. The structure of the electrolytic silver refining electrolytic cell is graphite as anode and stainless steel as cathode. The acid solution after dissolving the anode mud is used as electrolyte, and the current density is 200-400 A / m 2 , the electrolysis voltage is 1~5V, and the electrolysis time is 2~5h.

[0021] The polymetallic sulphur slag of the present invention is usually a by-product produced during the smelting process, and specifically includes:

[0022] The blast furnace slag generated by sintering-blast furnace smelting in the lead smelting process, the leached slag after acid leaching of zinc after roasting zinc concentrate (ZnS) in the zinc smelting process, or the hot melt filter cake after hot melt filtration in the zinc smelting process, etc.

[0023] The present invention primarily recovers sulfur slag based on the physical and chemical properties of the various phases. A sulfide salt is selected to dissolve elemental sulfur in the slag, achieving a phase transformation of the elemental sulfur. This process, when optimized, results in minimal subsequent metal loss. The low temperature (300-600°C) is selected primarily based on the decomposition temperature of mercuric sulfide and the evaporation temperature of mercury. The medium temperature (700-1000°C) is selected primarily because zinc sulfide begins to transform into an oxide in this temperature range. The subsequent acid leaching process allows for efficient zinc leaching, while lead and silver are largely unaffected under the same leaching conditions as zinc. Finally, lead is smelted at a high temperature (1100-1500°C). The refined crude lead is then used directly as the anode for electrolysis, where refined lead is deposited and separated at the cathode. The anode slime is then electrolyzed at the cathode via acid leaching, followed by silver recovery. To ensure consistent recovery of each metal in the aforementioned invention, the order of the steps must be maintained. If sulfur slag containing different metal components is being processed, the aforementioned steps may need to be appropriately omitted. If the order is disrupted, it will be difficult to achieve efficient partial recovery of sulfur and valuable metals, resulting in waste of resources.

[0024] The system of the present invention combines many existing technologies. The following are the ingenuity of the present invention:

[0025] 1. Closed-loop design of "step-by-step enrichment and directional transformation of physical phases"

[0026] Traditional processes are mostly focused on the recovery of a single metal, or the recovery rate is low due to impurities. This invention uses the chain connection of "slag → new raw materials" to achieve multi-metal step-by-step extraction:

[0027] The first step is sulfur recovery: chelation leaching is used to selectively separate sulfur, overcoming the separation difficulty of "sulfur and heavy metals mixed together" and replacing the high energy consumption and low efficiency of traditional hot melt / flotation;

[0028] The second step is mercury recovery: calcining the leaching residue, utilizing the volatility of mercury to separate the "mercury-removed residue + elemental mercury", converting harmful mercury into elemental mercury for recycling, and thus addressing environmental risks.

[0029] The third step is the recovery of zinc, silver and lead: the demercured slag is calcined again to achieve metal phase transformation → acid leaching is used to efficiently extract zinc → the separated slag is smelted to obtain silver-containing crude lead. By changing the phase through calcination, zinc, silver and lead are enriched in a targeted manner from the complex slag, breaking through the "separation interference when multiple metals coexist."

[0030] 2. “Targeted Adaptation” of Process Modules

[0031] In view of the complex composition and diverse phases of polymetallic sulphur slag, each link is equipped with exclusive technologies to solve the core pain points of traditional processes:

[0032] Sulfur extraction by chelation: This method avoids the high energy consumption and high pollution of traditional acid leaching / hot melting. It utilizes chelating agents to selectively dissolve sulfur and efficiently separate "encapsulated sulfur." ​​This method avoids the difficulty of traditional physical separation due to high viscosity and seeks breakthroughs through chemical leaching.

[0033] Calcination has a dual purpose: it not only volatilizes and recovers mercury to remove it, but also oxidizes and aggregates the metals through high temperature, providing high-quality raw materials for subsequent acid leaching and electrolysis, thus solving the dual problems of "mercury pollution" and "difficulty in extracting dispersed metals" in one fell swoop.

[0034] Graded application of electrolysis: Crude lead containing silver is first electrolyzed to obtain refined lead + anode mud, the anode mud is acid-dissolved and then electrolyzed to extract silver; graded electrolysis avoids co-precipitation of impurities, improves the purity of silver and lead, and breaks through the "mutual interference during co-electrolysis of multiple metals."

[0035] 3. Collaborative recycling of multiple metals across the entire value chain

[0036] Traditional processes often cause some elements, such as mercury and silver in sulfur slag, to be discarded or treated at low value due to interference between metals and process fragmentation. This invention incorporates all elements of sulfur, mercury, zinc, silver, and lead into the recovery system, achieving a win-win situation for both environmental protection and economy: mercury recovery eliminates some hazardous waste risks, and efficient recovery of sulfur, zinc, silver, and lead increases profits. Compared with traditional storage or single recovery, resource utilization is greatly improved;

[0037] Innovation in technology integration: It is not a simple splicing of existing technologies, but rather a series connection of modules of "leaching → calcination → acid leaching → electrolysis" to allow the "by-products" of the previous process to become the "raw materials" of the subsequent process, such as leaching residue → calcination raw materials, separation slag → silver-containing crude lead raw materials, forming a closed-loop logic of "zero waste slag". This is the "material flow closed-loop thinking" that is easily overlooked in general process design.

[0038] In general, the ingenuity of this invention lies in its "phase transformation as the core, chain connection as the path, and multi-element recovery as the goal", breaking through the traditional pain points of "difficulty in separating multi-metal mixtures, high cost of hazardous waste treatment, and low resource utilization", and integrating scattered technical modules into an efficient and coordinated recycling system. This design idea breaks away from the limitation of "single technology to solve a single problem" and embodies the systematic innovation of process design.

[0039] Different from the prior art, the present invention has the following beneficial effects:

[0040] (1) The present invention can efficiently separate the metal phase and elemental sulfur in the polymetallic sulfur slag, mainly by wet leaching the sulfur in the polymetallic sulfur slag and reconstructing the phase by pH control to precipitate and recover the elemental sulfur. The metal phases such as mercury sulfide, zinc sulfide, lead sulfide and silver sulfide that originally existed stably in the slag are not easily dissolved and lost, and can be retained in the slag for subsequent pyrometallurgical treatment.

[0041] (2) The present invention combines pyrometallurgy with the control of the conversion of valuable metals such as mercury, zinc, lead, and silver in the form of sulfides into corresponding metal elements or oxides, which are easily recovered by electrolysis after sufficient leaching by wet method. The valuable metals in the multi-metallic sulfur slag can be effectively enriched and separated, and the overall recovery rate of various metals can reach more than 90%. BRIEF DESCRIPTION OF THE DRAWINGS

[0042] In order to more clearly illustrate the technical solutions in the prior art of the present invention, the following are the drawings required for use in the embodiments or the description of the prior art. The drawings described below are only some embodiments of the present invention. For ordinary technicians in this field, other drawings can be obtained based on the structures shown in these drawings without paying any creative work.

[0043] Figure 1 This is a process flow chart for the graded recovery of sulfur and metals from polymetallic sulfur slag by combining the wet method and the pyrolysis method according to the present invention.

[0044] Figure 2 1 is the XRD pattern of the multi-metallic sulfur slag before and after the wet and active process combined in Example 1. The enriched slag treated by the method of the present invention in the figure corresponds to the mercury-free metal-enriched slag S3 in the present invention. DETAILED DESCRIPTION

[0045] The following will clearly and completely describe the technical solutions in the embodiments of the present invention in conjunction with the accompanying drawings. Obviously, the described embodiments are only part of the embodiments of the present invention, not all of the embodiments. Based on the embodiments of the present invention, all other embodiments obtained by ordinary technicians in this field without making any creative efforts are within the scope of protection of the present invention.

[0046] Furthermore, the technical solutions of the various embodiments of the present invention may be combined, but only if they are achievable by persons of ordinary skill in the art. If the combination of technical solutions contradicts or is unachievable, such combination shall be deemed non-existent and outside the scope of protection claimed by the present invention. Furthermore, the descriptions of 1 and 2 herein serve only to distinguish between different processes or substances.

[0047] When numerical ranges are given in the examples, it should be understood that unless otherwise specified herein, both endpoints of each numerical range and any value between the two endpoints may be used. Unless otherwise defined, all technical and scientific terms used herein are consistent with the prior art as understood by those skilled in the art and the description of the present invention. Any prior art methods, devices, and materials similar or equivalent to those described in the examples of the present invention may also be used to implement the present invention.

[0048] Unless otherwise specified, the experimental methods used in the following examples are conventional methods.

[0049] Unless otherwise specified, the materials and reagents used in the following examples can be obtained from commercial sources.

[0050] Example 1

[0051] The present invention provides a method for recovering sulfur and metals from polymetallic sulfur slag by combining a wet method and a pyrolysis method, comprising the following steps:

[0052] 1) The hot melt filter cake obtained after hot melt filtration during zinc smelting is simply crushed (less than 1 cm) and then added to a stirring tank. The slag has a sulfur content of 48.9%, a zinc content of 7.3%, a mercury content of 0.1%, a lead content of 1.4%, and a silver content of 0.07%. A chelating leaching agent (90% sodium sulfide and 10% ammonium sulfide) with a concentration of 15 wt% is prepared and mixed with polymetallic sulfur slag in a stirring tank at a liquid-solid ratio of 15 L / kg. The mixture is heated to 30° C. and stirred for 0.5 h at a stirring speed of 500 rpm / min to dissolve elemental sulfur. The mixture is filtered to achieve liquid-solid separation, and the leachate is collected and 10 wt% of a pH regulator (60% sulfuric acid and 40% hydrochloric acid) is added to adjust the pH to 8. The sodium polysulfide is controlled to precipitate elemental sulfur to achieve a sulfur purity of 98.8% and a recovery rate of 94.9%.

[0053] 2) The separated high-metal residue was dried and crushed, passed through a 100-mesh sieve, and then heated to 500° C. in an air atmosphere. After calcination for 3 hours, the liquid elemental mercury was cooled and recovered in the tail gas, achieving a mercury recovery rate of 93.1%.

[0054] 3) The mercury removal separation slag was crushed and passed through a 100-mesh sieve, then heated to 700° C. in an air atmosphere and calcined for 3 hours to obtain a mercury removal metal-enriched slag having a zinc content of 32.35%. The slag was dried, crushed, and passed through a 100-mesh sieve, and then leached at 80° C. for 3 hours with a 15 wt% acid solution (70% sulfuric acid and 30% hydrochloric acid). After liquid-solid separation, the zinc leaching solution was adjusted to a pH of 3 to separate and remove iron and arsenic as insoluble compounds. The temperature was then adjusted to 70° C. and 1.5 times the theoretical amount of zinc powder was added for replacement. Impurities such as copper and cadmium were separated and removed. An aluminum plate was used as a cathode and crude lead containing 0.5-1% silver was used as an anode. The solution after zinc powder replacement and removal of copper, cadmium and other impurities was used as an electrolyte. The electrolyte was electrolyzed at a current density of 500 A / m 2 The electrolysis voltage was 3V, the electrolysis time was 6h, and the elemental zinc was obtained by mechanical stripping at the cathode with a purity of 99.3% and a recovery rate of 97.8%.

[0055] 4) The dezincified metal-enriched slag was dried and crushed, passed through a 100-mesh sieve, and then heated to 1300°C in an air atmosphere and calcined for 5 hours to obtain silver-containing crude lead with a lead content of 96.54%. The lead content was 96.54% and used as the anode. Stainless steel was used as the cathode. A 10 wt% lead fluorosilicate solution was used as the electrolyte. The current density was 350 A / m 2 , the electrolysis voltage is 2V, the electrolysis time is 5h, the refined lead is mechanically stripped after being deposited on the cathode, the purity is 99.5%, and the recovery rate is 92.1%; the anode mud is dissolved in a 15wt% concentration acid solution (70% nitric acid, 30% sulfuric acid), graphite is used as the anode, stainless steel is used as the cathode, the acid solution after dissolving the anode mud is used as the electrolyte, and the filtrate after filtration is subjected to a current density of 300A / m 2 The electrolysis voltage was 3 V, the electrolysis time was 5 h, and elemental silver was obtained by mechanical stripping at the cathode with a purity of 99.1% and a recovery rate of 92.9%.

[0056] Example 2

[0057] The present invention provides a method for recovering sulfur and metals from polymetallic sulfur slag by combining a wet method and a pyrolysis method, comprising the following steps:

[0058] 1) Blast furnace slag generated by sintering-blast furnace smelting in a lead smelting process is simply crushed (less than 1 cm) and added to a stirring tank. The slag has a sulfur content of 54.9%, a zinc content of 8.3%, a mercury content of 0.3%, a lead content of 7.7%, and a silver content of 0.14%. A chelating leaching agent (80% sodium sulfide and 20% sodium hydroxide) with a concentration of 20 wt% is prepared and mixed with polymetallic sulfur slag in a stirring tank at a liquid-solid ratio of 20 L / kg. The mixture is heated to 50° C. and conventionally stirred for 2.0 hours at a stirring speed of 600 rpm / min to dissolve elemental sulfur. Liquid-solid separation is achieved by filtration. The collected leachate is then added with 15 wt% of a pH regulator (50% sulfuric acid and 50% hydrochloric acid) to adjust the pH to 7.5, thereby controlling the precipitation of elemental sulfur from sodium polysulfide to achieve a sulfur purity of 99.7% and a recovery rate of 96.9%.

[0059] 2) The separated high-metal residue was dried and crushed, passed through a 100-mesh sieve, and then heated to 600° C. in a pure oxygen atmosphere. After calcination for 5 hours, the liquid elemental mercury was cooled and recovered in the tail gas, achieving a mercury recovery rate of 95.4%.

[0060] 3) The mercury-removed separation slag was crushed and passed through a 100-mesh sieve, then heated to 800° C. in a pure oxygen atmosphere and calcined for 4 h to obtain a mercury-removed metal-enriched slag having a zinc content of 40.35%. The slag was dried, crushed, and passed through a 100-mesh sieve, and then leached at 100° C. with a 20 wt% acid solution (70% nitric acid, 30% hydrochloric acid) for 5 h. After liquid-solid separation, the zinc leaching solution was adjusted to a pH of 2 to separate and remove iron and arsenic as insoluble compounds. The temperature was then adjusted to 60° C. and zinc powder (100% of the theoretical amount) was added for replacement, followed by separation and removal of copper, cadmium and other impurities. An aluminum plate was used as the cathode, and crude lead containing 0.5-1% silver was used as the anode. The solution after zinc powder replacement and separation to remove copper, cadmium and other impurities was used as the electrolyte. The slag was electrolyzed at a current density of 550 A / m 2 The electrolysis voltage was 4V, the electrolysis time was 5h, and the elemental zinc was obtained by mechanical stripping at the cathode with a purity of 99.6% and a recovery rate of 99.4%.

[0061] 4) The dezincified metal-enriched slag was dried and crushed, passed through a 100-mesh sieve, and then heated to 1500°C in a pure oxygen atmosphere and calcined for 5 hours to obtain silver-containing crude lead with a lead content of 97.43%. The crude lead was used as the anode, stainless steel as the cathode, and 15 wt% lead fluorosilicate solution as the electrolyte. The current density was 400 A / m 2 , the electrolysis voltage is 3V, the electrolysis time is 8h, the refined lead is mechanically stripped after being deposited on the cathode, the purity is 99.7%, and the recovery rate is 95.9%; the anode mud is dissolved in a 20wt% concentration acid solution (70% nitric acid, 30% sulfuric acid), graphite is used as the anode, stainless steel is used as the cathode, the acid solution after dissolving the anode mud is used as the electrolyte, and the filtrate after filtration is subjected to a current density of 500A / m 2The electrolysis voltage was 5V, the electrolysis time was 8h, and the elemental silver was obtained by mechanical stripping at the cathode with a purity of 99.6% and a recovery rate of 97.7%.

[0062] Example 3

[0063] 1) After zinc concentrate (ZnS) is roasted and acid-leached to extract zinc from a zinc smelting process, the resulting slag is simply crushed (<1 cm) and added to a stirring tank. The slag has a S content of 65.6%, a zinc content of 11.3%, a mercury content of 0.4%, a lead content of 4.3%, and a silver content of 0.21%. A chelating leaching agent (70% potassium sulfide and 30% sodium sulfate) with a concentration of 15 wt% is prepared and mixed with polymetallic sulfur slag in a stirring tank at a liquid-solid ratio of 18 L / kg. The mixture is heated to 70° C. and conventionally stirred for 3.0 hours at a stirring speed of 300 rpm / min to dissolve elemental sulfur. Liquid-solid separation is achieved by filtration. The collected leachate is then added with 25 wt% of a pH regulator (100% sulfuric acid) to adjust the pH to 8.5, thereby controlling the precipitation of elemental sulfur from sodium polysulfide to achieve a sulfur purity of 97.9% and a recovery rate of 94.2%.

[0064] 2) The separated high-metal residue was dried and crushed, passed through a 100-mesh sieve, and then heated to 550° C. in a pure oxygen atmosphere. After calcination for 1 hour, the liquid elemental mercury was cooled and recovered in the tail gas, achieving a mercury recovery rate of 93.6%.

[0065] 3) The mercury-removed separation slag was crushed and passed through a 100-mesh sieve, and then calcined to 750° C. in a pure oxygen atmosphere to obtain a mercury-removed metal-enriched slag having a zinc content of 44.71%. The slag was dried, crushed, and passed through a 100-mesh sieve, and then leached at 90° C. for 2 h with a 30 wt% acid solution (100% nitric acid). After liquid-solid separation, the zinc leaching solution was adjusted to a pH of 3 to separate and remove iron and arsenic as insoluble compounds. The temperature was then adjusted to 65° C. and zinc powder (twice the theoretical amount) was added for replacement. Impurities such as copper and cadmium were separated and removed. An aluminum plate was used as a cathode, and crude lead containing 0.5-1% silver was used as an anode. The solution after zinc powder replacement and removal of impurities such as copper and cadmium was used as an electrolyte. The electrolyte was then used at a current density of 600 A / m 2 The electrolysis voltage was 5V, the electrolysis time was 8h, and the elemental zinc was obtained by mechanical stripping at the cathode with a purity of 99.2% and a recovery rate of 98.9%.

[0066] 4) The dezincified metal-enriched slag was dried and crushed, passed through a 100-mesh sieve, and then heated to 1400°C in a pure oxygen atmosphere and calcined for 5 hours to obtain silver-containing crude lead with a lead content of 94.41%. The lead content was 94.41% and used as the anode. Stainless steel was used as the cathode. A 20 wt% lead fluorosilicate solution was used as the electrolyte. The current density was 450 A / m 2, electrolysis voltage is 3V, electrolysis time is 3h, refined lead is mechanically stripped after cathode deposition, purity is 99.3%, recovery rate is 94.6%; anode mud is dissolved in 10wt% concentration acid solution (100% sulfuric acid), graphite is used as anode, stainless steel is used as cathode, the acid solution after dissolving the anode mud is used as electrolyte, the filtrate after filtration is at a current density of 350A / m 2 The electrolysis voltage was 4 V, the electrolysis time was 4 h, and the elemental silver was mechanically stripped at the cathode with a purity of 98.6% and a recovery rate of 95.4%.

[0067] Example 4

[0068] 1) The hot melt filter cake polymetallic sulfur slag obtained after hot melt filtration during zinc smelting is simply crushed (less than 1 cm) and added to a stirring tank. The slag contains 74.3% S, 10.4% zinc, 0.35% mercury, 6.1% lead, and 0.25% silver. A chelating leaching agent (70% sodium sulfide and 30% sodium sulfate) with a concentration of 20 wt% is prepared and mixed with the polymetallic sulfur slag in a stirring tank at a liquid-solid ratio of 20 L / kg. The mixture is heated to 80° C. and stirred for 5 hours at a stirring speed of 800 rpm / min to dissolve elemental sulfur. Liquid-solid separation is achieved by filtration. The leachate is collected, and 5 wt% of a pH regulator (30% sulfuric acid and 70% hydrochloric acid) is added to adjust the pH to 9. The sodium polysulfide is controlled to precipitate elemental sulfur to achieve a sulfur purity of 98.7% and a recovery rate of 90.8%.

[0069] 2) The separated high-metal residue was dried and crushed, passed through a 100-mesh sieve, and then heated to 500° C. in an air atmosphere. After calcination for 3 hours, the liquid elemental mercury was cooled and recovered in the tail gas, achieving a mercury recovery rate of 91.8%.

[0070] 3) The mercury-removed separation slag was crushed and passed through a 100-mesh sieve, then heated to 900° C. in an air atmosphere and calcined for 4 hours to obtain a mercury-removed metal-enriched slag having a zinc content of 46.42%. The slag was dried, crushed, and passed through a 100-mesh sieve, and then leached at 120° C. for 4 hours with a 25 wt% acid solution (30% nitric acid, 70% hydrochloric acid). After liquid-solid separation, the zinc leaching solution was adjusted to a pH of 1 to separate and remove iron and arsenic as insoluble compounds. The temperature was then adjusted to 80° C. and 1.8 times the theoretical amount of zinc powder was added to replace the slag. Impurities such as copper and cadmium were separated and removed. An aluminum plate was used as a cathode and crude lead containing 0.5-1% silver was used as an anode. The solution after zinc powder replacement and removal of copper, cadmium and other impurities was used as an electrolyte. The slag was electrolyzed at a current density of 550 A / m 2 The electrolysis voltage was 4 V, the electrolysis time was 6 h, and the elemental zinc was obtained by mechanical stripping at the cathode with a purity of 99.4% and a recovery rate of 97.3%.

[0071] 4) The dezincified metal-enriched slag was dried and crushed, passed through a 100-mesh sieve, and then heated to 1200°C in an air atmosphere and calcined for 5 hours to obtain silver-containing crude lead with a lead content of 96.37%. The lead content was 96.37% and used as the anode. Stainless steel was used as the cathode. A 15 wt% lead fluorosilicate solution was used as the electrolyte. The current density was 150 A / m 2 , electrolysis voltage is 3V, electrolysis time is 5h, refined lead is mechanically stripped after cathode deposition, purity is 99.4%, recovery rate is 95.7%; anode mud is dissolved in 20wt% concentration acid solution (nitric acid 100%), graphite is used as anode, stainless steel is used as cathode, the acid solution after dissolving the anode mud is used as electrolyte, the filtrate after filtration is at a current density of 300A / m 2 The electrolysis voltage was 3V, the electrolysis time was 5h, and the elemental silver was obtained by mechanical stripping at the cathode with a purity of 99.1% and a recovery rate of 96.8%.

[0072] Comparative Example 1

[0073] The same polymetallic sulfur slag as in Example 4 of the present invention was treated by a conventional roasting-leaching process and compared with the effect of the present invention, which includes the following steps:

[0074] The polymetallic sulfur slag containing 74.3% S, 10.4% zinc, 0.35% mercury, 6.1% lead and 0.25% silver was dried and crushed to pass through a 100-mesh sieve, and then heated to 500°C in a pure oxygen atmosphere and calcined for 3 hours. The flue gas then entered the acid-making system to produce sulfuric acid, with a S recovery rate of only 83.4%. Mercury entering the flue gas was difficult to recover, with a recovery rate of 0%. The slag was dried and crushed to pass through a 100-mesh sieve, and then leached at 120°C with a 25wt% acid solution (30% nitric acid, 70% hydrochloric acid) for 4 hours. After liquid-solid separation, the zinc leachate was adjusted to pH 1 to separate and remove iron and arsenic into insoluble compounds. The temperature was then adjusted to 80°C and zinc powder twice the theoretical amount was added to replace the iron and arsenic. After the copper, cadmium and other impurities were separated and removed, an aluminum plate was used as the cathode and crude lead containing 0.5-1% silver was used as the anode. After the zinc powder was replaced and the copper, cadmium and other impurities were separated and removed, the solution became the electrolyte. The electrolyte was electrolytically tested at a current density of 600A / m 2 The electrolysis voltage was 4V and the electrolysis time was 6h. Elemental zinc was mechanically stripped at the cathode with a purity of 34.1% and a recovery rate of 10.2%. The lead content in the leaching separation residue was 35.7% and the silver content was 7.8%, which was difficult to recover.

[0075] The above-mentioned traditional roasting-leaching process does not have any subsequent further treatment of lead and silver after the leaching liquid-solid separation; due to the low zinc recovery rate caused by the previous process, most of the zinc enters the leaching residue, which forms a lead-silver-zinc composite slag after high-temperature smelting. The subsequent electrolysis step will also make it difficult to effectively separate them.

[0076] In summary, the present invention provides a method for separating and recovering valuable metals by regulating sulfur phase reconstruction through wet method and combining it with pyrolysis to coordinate zinc, lead and silver multi-metal phase transformation. The method is easy to operate and has good treatment effect, effectively realizes the separation of metal and sulfur, and simultaneously realizes the controllable transformation and refining of mercury, zinc, lead and silver valuable metal phases, converts sulfur element into elemental sulfur for separation, and comprehensively recovers valuable metals and sulfur resources in hazardous wastes, with significant economic benefits.

[0077] The technical solution of the present invention is not limited to the above-mentioned specific embodiments. Any technical variations made according to the technical solution of the present invention fall within the protection scope of the present invention.

Claims

1. A method for recovering sulfur and metals from polymetallic sulfur slag by combining wet and pyrolysis, characterized in that: The following steps are involved: S1, mixing the polymetallic sulfur slag with a chelating leaching agent, performing a leaching reaction, and obtaining a sulfur leachate and a polymetallic residue after solid-liquid separation, wherein the sulfur leachate is reconstituted into sulfur by adjusting the pH; S2, drying and crushing the multi-metallic residue, calcining to obtain a mercury-removed separation slag, and cooling and collecting the tail gas during the calcination process to obtain elemental mercury; S3, crushing and calcining the demercured separation slag to obtain demercured metal-enriched slag; S4, drying and crushing the demercured metal-enriched slag, performing a leaching reaction with an acid solution, obtaining a zinc leachate and dezincified metal-enriched slag after solid-liquid separation, and then neutralizing and replacing the zinc leachate, purifying and electrolyzing it to obtain elemental zinc; S5, drying and crushing the dezincified metal-enriched slag, and then calcining it to obtain silver-containing crude lead, which is then refined by electrolytic refining, wherein the anode mud is acid-dissolved and then electrolyzed to obtain elemental silver.

2. The method for recovering sulfur and metals from polymetallic sulfur slag by combining wet and pyrometallurgical methods according to claim 1, characterized in that: The chelating leaching agent in S1 includes: at least one of sodium sulfide, potassium sulfide, ammonium sulfide, sodium sulfate, and sodium hydroxide; the concentration of the chelating leaching agent is 10-20wt%, the liquid-solid ratio is 10-20L / kg, the leaching time is 0.5-4h, the temperature is 25-80°C, and the stirring speed is 100-800rpm / min.

3. The method for recovering sulfur and metals from polymetallic sulfur slag by combining wet and pyrometallurgical methods according to claim 1 or 2, characterized in that: In the step S1, the pH is adjusted to a range of 7.5 to 12.5, and the pH regulator is at least one of sulfuric acid and hydrochloric acid, with a concentration of 5 to 30 wt%.

4. The method for recovering sulfur and metals from polymetallic sulfur slag by combining wet and pyrometallurgical methods according to claim 1, characterized in that: In S2, the multi-metal residue is dried, crushed, passed through a 100-mesh sieve, and then calcined. The calcination atmosphere is at least one of air and oxygen, the temperature range is 300-600° C., and the calcination time is 1-5 hours.

5. The method for recovering sulfur and metals from polymetallic sulfur slag by combining wet and pyrometallurgical methods according to claim 1 or 4, characterized in that: In S3, the mercury removal separation slag is crushed and passed through a 100-mesh sieve and then calcined. The calcination atmosphere is at least one of air and oxygen. The calcination temperature ranges from 700 to 1000° C. and the calcination time is 1 to 5 hours.

6. The method for recovering sulfur and metals from polymetallic sulfur slag by combining wet and pyrometallurgical methods according to claim 1, characterized in that: In S4, the demercured metal-enriched slag is dried, crushed, passed through a 100-mesh sieve, and then used for acid leaching. The acid leaching temperature range is 60-120° C., the acid solution is at least one of sulfuric acid, hydrochloric acid, and nitric acid, the concentration is 10-30 wt%, and the leaching time is 1-5 hours.

7. The method for recovering sulfur and metals from polymetallic sulfur slag by combining wet and pyrometallurgical methods according to claim 1 or 6, characterized in that: In the step S4, the zinc leaching solution is adjusted to a pH of 1 to 4, so that iron and arsenic are separated and removed by forming insoluble compounds. Subsequently, the temperature is adjusted to 60 to 85°C and 1.5 to 2 times the theoretical amount of zinc powder is added to replace the copper, cadmium and other impurities. The electrolytic zinc smelting electrolytic cell structure is composed of an aluminum plate as a cathode and a crude lead containing 0.5 to 1% silver as an anode. The solution after zinc powder replacement and separation to remove copper, cadmium and other impurities is used as the electrolyte. The current density is 400 to 600 A / m 2 , the electrolysis voltage is 1~4V, and the electrolysis time is 3~10h.

8. The method for recovering sulfur and metals from polymetallic sulfur slag by combining wet and pyrometallurgical methods according to claim 1, characterized in that: In S5, the dezincified metal-enriched slag is dried, crushed, and passed through a 100-mesh sieve, and then calcined. The calcination atmosphere is oxygen, and the temperature is 1100-1500° C. The electrolytic lead smelting cell structure is silver-containing crude lead as the anode, stainless steel as the cathode, 10-20wt% lead fluoride silicate solution as the electrolyte, and the current density is 300-500A / m 2 , the electrolysis voltage is 1~3V, and the electrolysis time is 3~10h.

9. The method for recovering sulfur and metals from polymetallic sulfur slag by combining wet and pyrometallurgical methods according to claim 1 or 8, characterized in that: The acid solution for dissolving the anode mud in S5 includes: at least one of sulfuric acid and nitric acid, with a concentration of 10-20 wt%. The structure of the electrolytic silver refining electrolytic cell is graphite as anode and stainless steel as cathode. The acid solution after dissolving the anode mud is used as electrolyte, and the current density is 200-400 A / m 2 , the electrolysis voltage is 1~5V, and the electrolysis time is 3~10h.

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