Method for recovering and smelting rare and precious metals from acid sludge
By employing green oxidants and agglomeration flotation, rare and precious metals are efficiently extracted from smelting acid sludge, solving the environmental and energy consumption problems of existing technologies. This achieves high selenium recovery rate and efficient silver flotation, while reducing costs.
Patent Information
- Authority / Receiving Office
- CN · China
- Patent Type
- Applications(China)
- Current Assignee / Owner
- HUNAN ZHUYE ENVIRONMENT TECH CO LTD
- Filing Date
- 2026-05-20
- Publication Date
- 2026-06-30
Smart Images
Figure CN122303612A_ABST
Abstract
Description
Technical Field
[0001] This invention belongs to the field of metallurgical solid waste resource utilization technology, specifically relating to a method for recovering rare and precious metals from smelting acid sludge. Background Technology
[0002] During the acid production process in lead-zinc smelting systems, a large amount of mercury-containing acid sludge is generated. This sludge is typically removed at high temperatures using a roasting method to obtain crude mercury. The remaining smelting acid sludge after mercury removal contains other rare and precious metals such as selenium and silver. Utilizing this smelting acid sludge as a resource is essential; extracting selenium, silver, and other rare and precious metals can generate additional economic benefits.
[0003] Currently, the treatment method for smelting acid sludge involves selectively leaching selenium first, followed by the extraction of rare and precious metals such as silver. In actual engineering, incomplete oxidation of the acid sludge during roasting often results in selenium in the smelting acid sludge existing in the form of CaSe. To solve this problem, traditional processes generally use sulfuric acid-nitric acid oxidation leaching or sodium chlorate leaching to oxidize Se in CaSe into selenate or selenite. Chinese patent CN202210114421.1 discloses a hydrometallurgical method for treating high-arsenic, low-mercury selenate acid sludge, in which sodium chlorate is used as an oxidant in a hydrochloric acid system to obtain a selenium-containing solution. However, this step may generate excess toxic chlorine gas. Although the sulfuric acid-nitric acid oxidation leaching system can effectively leach Se, the large amount of nitrogen oxides generated during the reaction pollutes the environment. Therefore, the oxidative leaching process of smelting acid sludge needs to be optimized for environmental protection.
[0004] Gypsum slag, after oxidation and leaching, generally contains precious metals such as silver, but there is currently no research on directly extracting silver from it. Current research and patents involve extracting valuable metals such as copper, iron, and zinc from gypsum slag, primarily using reduction roasting as the core process to induce phase transformation of the target metal elements, facilitating further extraction in subsequent processes. This is especially beneficial for fine-grained target minerals, as the roasting process can increase the crystal size, thus aiding downstream beneficiation and enrichment processes. Chinese patent CN 201910440146.0 discloses a method for enhanced recovery of valuable metals from copper smelting slag using gypsum slag. This method involves reduction roasting at temperatures above 1000℃ and controlling the cooling rate to achieve crystal growth of the target mineral, thereby enabling downstream beneficiation. However, reduction roasting technology itself is an energy-intensive process, potentially leading to high costs.
[0005] In conclusion, the rare and precious metal recovery process of conventional smelting acid sludge still needs to be optimized. Summary of the Invention
[0006] The purpose of this invention is to optimize some problems existing in the rare and precious metal extraction process of smelting acid sludge, and to form a new rare and precious metal extraction process. Existing processes for extracting selenium from calcined sand traditionally use a sulfuric acid-nitric acid system, but this process involves environmental issues and the consumption of large amounts of reagents. This invention uses the green and clean oxidant hydrogen peroxide, avoiding the problem of harmful gases during the oxidative leaching process and preventing the back dissolution problem caused by excessive nitric acid during selenium reduction. On the other hand, the gypsum slag produced after oxidative leaching contains a considerable amount of rare and precious metals, such as silver. However, existing research has not mentioned the selective extraction of silver from gypsum slag, and the silver-containing crystal grain size in metallurgical slag is generally less than 10 micrometers, which is not conducive to flotation and gravity separation processes. This invention proposes for the first time the selective extraction of silver from gypsum slag through a flotation process, and solves the problem of difficulty in flotation due to the small crystal grain size of silver-containing minerals.
[0007] Specifically, the present invention provides a method for recovering rare and precious metals from smelting acid sludge, the method comprising: (1) The smelting acid mud was oxidized and leached using a sulfuric acid-hydrogen peroxide leaching system to obtain a selenium-containing solution and gypsum slag. The amount of hydrogen peroxide added was 1-5 wt%, and the amount of sulfuric acid added was 300-500 g / L. The liquid-solid ratio was 5-10:1. The Hg in the smelting acid mud was basically removed and contained a certain amount of Se, Ag and other precious metals. Due to the incomplete oxidation during the mercury removal process, the Se in the calcined sand mainly existed in the form of CaSe. The main phases of the smelting acid mud were CaO and CaSe. The liquid-solid ratio was adjusted by adding water. During the reaction, under the premise of meeting the above parameters, the leaching temperature of the system was 25-40℃ and no additional heating was required. (2) The selenium-containing solution is reduced by a reducing agent to obtain crude selenium. Since the selenium-containing solution in step (1) is strongly acidic, it can be directly reduced by a reducing agent to obtain crude selenium. The selenium recovery rate reaches more than 89%, and the crude selenium grade reaches more than 95%. (3) The gypsum residue obtained in step (1) is pre-sulfurized, and then the solid and liquid are separated to obtain the sulfurized gypsum residue; (4) The silver and other precious metals in the sulfided gypsum slag obtained in step (3) are enriched by agglomeration flotation, including: primary roughing, primary scavenging, and secondary scavenging. The primary roughing yields silver concentrate and primary tailings. The primary tailings are processed by primary scavenging to obtain primary middlings and secondary tailings. The secondary tailings are processed by secondary scavenging to obtain secondary middlings and secondary tailings. The primary middlings are returned to the next primary roughing, and the secondary middlings are returned to the next primary scavenging. The secondary tailings are mainly gypsum after the rare and precious metals have been extracted and can be used as building materials. In the first roughing, first scavenging, and second scavenging processes, dispersants, agglomerants, collectors, and foaming agents are added sequentially, respectively. The agglomerants are used to selectively agglomerate fine silver-containing crystals and other fine-grained precious metal particles.
[0008] The principles of oxidative leaching and selenium reduction in smelting acid mud are as follows: CaSe+3H2O2+H2SO4=CaSO4+3H2O+H2SeO3(1) H2SeO3+2Na2SO3=2Na2SO4+Se+H2O(2) As a preferred option, the above-mentioned method for recovering rare and precious metals from smelting acid mud includes, in step (1), the following steps are taken before oxidative leaching: crushing and grinding the smelting acid mud to below 200 mesh. Generally, the blocky smelting acid mud is crushed and ground to 200 mesh, which can increase the specific surface area, strengthen the contact between the leaching agent and the mineral, improve the selenium leaching rate, avoid incomplete leaching of large pieces of material, and ensure that the leaching reaction is sufficient and stable.
[0009] As a preferred option, in the above-mentioned method for recovering rare and precious metals from smelting sludge, the leaching time in step (1) is 2-6 hours. Controlling the leaching time to 2-6 hours can ensure that selenium is fully leached out, while avoiding excessive time that would lead to increased energy consumption and decreased efficiency, thus balancing the leaching rate and production cost.
[0010] As a preferred option, in the above-mentioned method for recovering rare and precious metals from smelting acid sludge, in step (2), the selenium-containing solution is heated and reduced with a reducing agent at 60-80℃. This temperature can increase the reduction rate and promote the rapid precipitation of crude selenium. At the same time, it can effectively decompose residual hydrogen peroxide, eliminate interference from oxidants, avoid selenium back dissolution, and improve the grade and recovery rate of crude selenium.
[0011] As a preferred embodiment, in the above-mentioned method for recovering rare and precious metals from smelting acid sludge, the reducing agent in step (2) is hydrazine hydrate and / or sodium sulfite. Hydrazine hydrate and sodium sulfite have strong reducing power and good selectivity, and can efficiently reduce selenite to elemental selenium; both are widely available, low in cost, and their byproducts are easy to handle, making them suitable for industrial application.
[0012] As a preferred option, in the above-mentioned method for recovering rare and precious metals from smelting sludge, step (3) involves pre-sulfurization using sodium sulfide solution for 2-10 minutes, with a sodium sulfide dosage of 500-1000 g / t gypsum slag. Sodium sulfide can modify the surface of silver-containing minerals in gypsum slag through sulfidation, enhancing hydrophobicity and significantly improving the floatability of silver minerals. With the dosage and time controlled within this range, sulfidation is sufficient and not excessive, resulting in low reagent consumption and stable effects.
[0013] As a preferred option, in the above-mentioned method for recovering rare and precious metals from smelting sludge, the slurry concentration of gypsum slag in step (3) is 20-30%. At this concentration, the slurry has good fluidity, the reagents are evenly dispersed, and the sulfidation reaction is sufficient; if the concentration is too low, the efficiency is low, and if it is too high, it is easy to agglomerate. 20-30% takes into account both the reaction efficiency and the subsequent flotation indicators.
[0014] As a preferred embodiment, in the above-mentioned method for recovering rare and precious metals from smelting sludge, the dispersant in step (4) is sodium hexametaphosphate. Sodium hexametaphosphate can effectively disperse fine gypsum gangue, reduce gangue agglomeration and entrainment, reduce interference with silver mineral flotation, and improve the grade of silver concentrate.
[0015] As a preferred embodiment, in the above-mentioned method for recovering rare and precious metals from smelting sludge, the agglomerating agent in step (4) is emulsified kerosene. Emulsified kerosene has a selective bridging effect on hydrophobic particles in the slurry, but cannot form a surface oil film on particles such as CaSO4, which have a strongly hydrophilic surface. By using sulfidation to replace the slurry, Ag-containing particles in the gypsum slag are converted into hydrophobic Ag2S crystals as much as possible, which can then be selectively agglomerated by the agglomerating agent and captured by the collector.
[0016] As a preferred embodiment, in the above-mentioned method for recovering rare and precious metals from smelting sludge, in step (4), the composite collector is a mixture of black powder and xanthate, with a mass ratio of 3-5:1. This ratio combines the advantages of good selectivity of black powder and strong collecting ability of xanthate, resulting in high collection efficiency and good selectivity for silver sulfide minerals, which can significantly improve the grade and recovery rate of silver concentrate.
[0017] As a preferred option, in the above-mentioned method for recovering rare and precious metals from smelting acid sludge, the foaming agent in step (4) is pine oil. Pine oil foam is stable and moderately brittle, and can produce uniform and moderate foam, which is conducive to the floating of silver minerals. The foam is not sticky and is easy to defoam, which has little impact on subsequent operations.
[0018] As a preferred option, in the above-mentioned method for recovering rare and precious metals from smelting acid sludge, step (4) involves a primary roughing process: The pulp concentration is 20-30%; The dosage of dispersant is 500-1000 g / t, and the reaction time is optionally 5-10 min; The amount of agglomerant added is 20-50 g / t, and the reaction time is optionally 15-30 min; The dosage of the compound collector is 300-600 g / t, and the reaction time is optionally 4-8 min; The dosage of foaming agent is 20-50 g / t, and the reaction time is optionally 1-3 min; The time for flotation and bubble scraping in a single roughing process at natural pH and room temperature can be arbitrarily set to 4-6 minutes.
[0019] With this parameter combination, the dispersion, aggregation, collection, and foaming steps are fully and synergistically performed. The operation is simple at natural pH and room temperature, and it can efficiently recover easily floating silver minerals to obtain high-grade silver concentrate.
[0020] As a preferred option, in the above-mentioned method for recovering rare and precious metals from smelting acid sludge, step (4) involves the first scan: The dosage of dispersant is 200-600 g / t, and the reaction time is optionally 2-7 min; The amount of agglomerant added is 10-30 g / t, and the reaction time is optionally 10-20 min; The amount of compound collector added is 200-400 g / t, and the reaction time is optionally 3-6 min; The dosage of foaming agent is 10-20 g / t, and the reaction time is optionally 1-3 min; The time for the first scavenging flotation at natural pH and room temperature can be arbitrarily set to 2-4 minutes.
[0021] By reducing reagent dosage and shortening the time, medium-floatable silver minerals in roughing tailings can be further recovered at low cost, reducing silver loss and increasing the overall recovery rate.
[0022] As a preferred option, in the above-mentioned method for recovering rare and precious metals from smelting acid sludge, step (4) involves a second scan: The dosage of dispersant is 100-300 g / t, and the reaction time is optionally 4-6 min; The amount of agglomerant added is 10-20 g / t, and the reaction time is optionally 10-20 min; The amount of compound collector added is 100-200 g / t, and the reaction time is optionally 4-6 min; The dosage of foaming agent is 5-10 g / t, and the reaction time is optionally 1-3 min; The second scavenging process, at natural pH and room temperature, can be performed for 2-4 minutes of flotation skimming.
[0023] Low reagent dosage and mild flotation conditions can effectively recover fine and difficult-to-float silver particles, while reducing gangue flotation, improving overall recovery rate and ensuring concentrate grade.
[0024] The advantages of this invention are: (1) This invention utilizes the resources of smelting acid mud to achieve efficient recovery of rare and precious metals.
[0025] (2) The oxidative leaching process in this invention is different from the traditional sulfuric acid-nitric acid leaching system. It uses green and clean hydrogen peroxide as an oxidant, does not introduce other interfering elements, avoids the problem of resolution during selenium reduction, thereby reducing the amount of reducing agent. Moreover, the leaching temperature is lower than that of the traditional oxidative leaching process, thereby reducing energy consumption and cost. (3) The reduction roasting process can promote the growth of the target mineral grain size and improve the floatability of the target mineral. However, the high-temperature roasting process is a high-energy-consuming and high-cost technical means. In order to reduce the cost of obtaining fine-grained rare and precious metal minerals, this invention avoids the roasting process and directly uses an agglomerating agent to selectively agglomerate the fine-grained rare and precious metal particles, thereby improving floatability and greatly reducing the cost of rare and precious metal extraction.
[0026] The above description is merely an overview of the technical solution of the present invention. In order to better understand the technical means of the present invention and to implement it in accordance with the contents of the specification, and to make the above and other objects, features and advantages of the present invention more apparent and understandable, specific embodiments of the present invention are described below. Attached Figure Description
[0027] Figure 1 A process route for extracting selenium from smelting acid mud; Figure 2 A process for extracting silver from gypsum residue; Figure 3 Comparative Example 1: SEM image of gypsum residue - containing silver crystals; Figure 4 Comparative Example 1: SEM images of the reduction product after reduction calcination, showing nano-Ag crystals. Detailed Implementation
[0028] The following description provides numerous specific details to offer a more thorough understanding of the technical solutions provided by this invention. However, it will be apparent to those skilled in the art that the technical solutions provided by this invention can be implemented without one or more of these details.
[0029] In the embodiments and comparative examples of this invention, the devices and raw materials used are commercially available.
[0030] Each embodiment provides a method for recovering rare and precious metals from smelting acid sludge; the process is described in detail below. Figure 1 , Figure 2 .
[0031] Example 1 120g of smelting acid mud was subjected to oxidative leaching at a liquid-to-solid ratio of 5:1, a leaching temperature of 30℃, a sulfuric acid concentration of 500g / L, a hydrogen peroxide concentration of 3%, and a leaching time of 4h. After leaching, 104.4g of gypsum residue and a selenium-containing leachate were obtained. The selenium-containing solution was then heated, and sodium sulfite was added to the selenium-containing leachate at 70℃ for reduction, yielding 18.99g of crude selenium with a grade of 95.64% and a recovery rate of 90.56%.
[0032] The gypsum slag was then sulfided for 3 minutes, with a sodium sulfide dosage of 600 g / t and a pulp concentration of 25%. After sulfidation, solid-liquid separation and drying were performed to obtain sulfidated gypsum slag. 100 g of the sulfidated gypsum slag was taken for agglomeration flotation. In the first roughing process, the pulp concentration was adjusted to 25%, and under natural pH conditions, 800 g / t of sodium hexametaphosphate was added and reacted for 10 minutes; 20 g / t of emulsified kerosene was added and reacted for 20 minutes; then 500 g / t of a composite collector was added, in which the ratio of dithiophosphate to xanthate was 4:1, and reacted for 6 minutes; finally, 20 g / t of pine oil was added and reacted for 2 minutes, and flotation was carried out for 5 minutes at natural pH and room temperature.
[0033] The tailings obtained from the first roughing process are then subjected to two more rounds of scavenging. The first round of scavenging involves adding 200 g / t of dispersant sodium hexametaphosphate and reacting for 4 minutes; adding 10 g / t of emulsified kerosene and reacting for 10 minutes; then adding 200 g / t of compound collector and reacting for 5 minutes; finally, adding 10 g / t of pine oil for a foaming reaction for 2 minutes, and then flotating at natural pH and room temperature for 3 minutes. This round of scavenging yields the first batch of middlings and tailings. The middlings will be returned to the next roughing step, while the tailings will require further scavenging.
[0034] The second round of scavenging requires the addition of 100 g / t of dispersant sodium hexametaphosphate and a reaction time of 5 min; the addition of 10 g / t of emulsified kerosene and a reaction time of 15 min; followed by the addition of 100 g / t of compound collector and a reaction time of 5 min; finally, the addition of 5 g / t of pine oil for a foaming reaction of 2 min, and flotation at natural pH and room temperature for 3 min; this round of scavenging yields a second batch of middlings and tailings, which will be returned to the first scavenging step of the next round. The results are shown in Table 1-2.
[0035] Table 1-1 Main elemental composition of mercury-removed roasted sand and sulfide gypsum residue
[0036] Table 1-2 Ag grade and recovery rate of mineral samples after agglomeration flotation
[0037] Example 2 120g of smelting acid mud was subjected to oxidative leaching at a liquid-to-solid ratio of 5:1, a leaching temperature of 30℃, a sulfuric acid concentration of 500g / L, a hydrogen peroxide concentration of 3%, and a leaching time of 4h. After leaching, 103.8g of gypsum residue and a selenium-containing leachate were obtained. The selenium-containing solution was then heated, and sodium sulfite was added to the selenium-containing leachate at 70℃ for reduction, yielding 21.82g of crude selenium with a grade of 96.12% and a recovery rate of 92.33%.
[0038] The gypsum slag was then sulfided for 8 minutes with a sodium sulfide dosage of 700 g / t and a pulp concentration of 30%. After sulfidation, solid-liquid separation and drying were performed to obtain sulfidated gypsum slag. 100 g of the sulfidated gypsum slag was taken for agglomeration flotation. In the first roughing process, the pulp concentration was adjusted to 25%, and under natural pH conditions, 900 g / t of sodium hexametaphosphate was added and reacted for 10 minutes; 30 g / t of emulsified kerosene was added and reacted for 25 minutes; then 600 g / t of a compound collector was added, in which the ratio of black reagent to xanthate was 4:1; finally, 20 g / t of pine oil was added and reacted for 2 minutes, followed by flotation for 5 minutes at natural pH and room temperature.
[0039] The tailings obtained from the first roughing process are then subjected to two more rounds of scavenging. The first round of scavenging involves adding 300 g / t of dispersant sodium hexametaphosphate and reacting for 5 minutes; adding 15 g / t of emulsified kerosene and reacting for 10 minutes; then adding 300 g / t of compound collector and reacting for 5 minutes; finally, adding 10 g / t of pine oil for a foaming reaction for 2 minutes and flotation at natural pH and room temperature for 3 minutes. This round of scavenging yields the first batch of middlings and tailings. The middlings will be returned to the next roughing process, while the tailings will require further scavenging.
[0040] The second round of scavenging requires the addition of 150 g / t of dispersant sodium hexametaphosphate and a reaction time of 5 min; the addition of 15 g / t of emulsified kerosene and a reaction time of 10 min; followed by the addition of 100 g / t of compound collector and a reaction time of 5 min; finally, the addition of 5 g / t of pine oil for a foaming reaction of 2 min and flotation at natural pH and room temperature for 3 min; this round of scavenging yields a second batch of middlings and tailings, which will be returned to the first scavenging step of the next round. The results are shown in Table 2-2.
[0041] Table 2-1 Main elemental composition of mercury-removed roasted sand and sulfide gypsum residue
[0042] Table 2-2 Ag grade and recovery rate of mineral samples after agglomeration flotation
[0043] Example 3 120g of smelting acid mud was subjected to oxidative leaching at a liquid-to-solid ratio of 5:1, a leaching temperature of 30℃, a sulfuric acid concentration of 450g / L, a hydrogen peroxide concentration of 2.5%, and a leaching time of 6h. After leaching, 102.9g of gypsum residue and a selenium-containing leachate were obtained. The selenium-containing solution was then heated, and sodium sulfite was added to the selenium-containing leachate at 70℃ for reduction, yielding 19.72g of crude selenium with a grade of 95.12% and a recovery rate of 89.23%.
[0044] The gypsum slag was then sulfided for 3 minutes, with a sodium sulfide dosage of 550 g / t and a pulp concentration of 25%. After sulfidation, solid-liquid separation and drying were performed to obtain sulfidated gypsum slag. 100 g of the sulfidated gypsum slag was taken for agglomeration flotation. In the first roughing process, the pulp concentration was adjusted to 25%, and under natural pH conditions, 1000 g / t of sodium hexametaphosphate was added and reacted for 10 minutes; 30 g / t of emulsified kerosene was added and reacted for 25 minutes; then 600 g / t of a compound collector was added, in which the ratio of black reagent to xanthate was 4:1; finally, 40 g / t of pine oil was added, reacted for 2 minutes, and flotation was carried out for 5 minutes at natural pH and room temperature.
[0045] The tailings obtained from the first roughing process are then subjected to two more rounds of scavenging. The first round of scavenging involves adding 300 g / t of dispersant sodium hexametaphosphate and reacting for 5 minutes; adding 20 g / t of emulsified kerosene and reacting for 10 minutes; then adding 300 g / t of compound collector and reacting for 5 minutes; finally, adding 10 g / t of pine oil for a foaming reaction for 2 minutes and flotation at natural pH and room temperature for 3 minutes. This round of scavenging yields the first batch of middlings and tailings. The middlings will be returned to the next roughing step, while the tailings will require further scavenging.
[0046] The second round of scavenging requires the addition of 100 g / t of dispersant sodium hexametaphosphate and a reaction time of 5 min; the addition of 10 g / t of emulsified kerosene and a reaction time of 15 min; followed by the addition of 100 g / t of compound collector and a reaction time of 5 min; finally, the addition of 5 g / t of pine oil for a foaming reaction of 2 min and flotation at natural pH and room temperature for 3 min; this round of scavenging yields a second batch of middlings and tailings, which will be returned to the first scavenging step of the next round. The results are shown in Table 3-2.
[0047] Table 3-1 Main elemental composition of mercury-removed roasted sand and sulfide gypsum residue
[0048] Table 3-2 Ag grade and recovery rate of mineral samples after agglomeration flotation
[0049] Because CaSO4 has the effect of encapsulating or connecting silver-containing grains, it is not conducive to the flotation of Ag ore. Although adding dispersants can reduce these negative effects to a certain extent, it is still metallurgical slag after all. Almost all the grains and gangue particles are very small. Compared with the beneficiation results of natural ore flotation, the indicators will fluctuate relatively. Therefore, the Ag recovery rate of each embodiment fluctuates within a certain range.
[0050] Comparative Example 1 10g of gypsum slag and 0.3g of coke were thoroughly mixed. The uniformly mixed material was then loaded into a corundum boat and placed in the constant-temperature zone of a tubular furnace. Under continuous nitrogen purging, the temperature was raised to 800℃ and then maintained at this temperature for 2 hours for reduction roasting. During this process, the coke acts as a reducing agent, reacting with the oxidizing components in the system while suppressing the influence of the oxidizing atmosphere and promoting the phase transformation of the target components. After roasting, the furnace was cooled to room temperature under nitrogen protection to finally obtain the reduction roasted product.
[0051] Figure 3 The SEM images show that the gypsum residue before reduction mainly consists of flaky calcium sulfate, with Ag-containing grains intercalated or encapsulated within the gypsum. The Ag is also clearly observed to exhibit an AgCl morphology (micron-sized). After reduction and calcination... Figure 4 The SEM image shows Ag elemental deposits on the CaS surface after reduction roasting. It can be clearly observed that almost all Ag elemental particles are smaller than 1 μm. This reduces the size of the target silver crystals in the reduction product and further reduces floatability. Therefore, reduction roasting is not conducive to the recovery of Ag.
[0052] Comparative Example 2 The main difference from the previous example is that emulsified kerosene is not added, and some parameters are slightly different.
[0053] 120g of smelting acid mud was subjected to oxidative leaching at a liquid-to-solid ratio of 5:1, a leaching temperature of 30℃, a sulfuric acid concentration of 500g / L, a hydrogen peroxide concentration of 3%, and a leaching time of 4h. After leaching, 103.1g of gypsum residue and a selenium-containing leachate were obtained. The selenium-containing solution was then heated, and sodium sulfite was added to the selenium-containing leachate at 70℃ for reduction, yielding 21.82g of crude selenium with a grade of 96.12% and a recovery rate of 92.33%.
[0054] The gypsum slag was then sulfided for 8 minutes, with a sodium sulfide dosage of 700 g / t and a pulp concentration of 30%. After sulfidation, solid-liquid separation and drying were performed to obtain sulfidated gypsum slag. 100 g of the sulfidated gypsum slag was taken for flotation. In the first roughing process, the pulp concentration was adjusted to 25%, and under natural pH conditions, 900 g / t of sodium hexametaphosphate was added and reacted for 10 minutes. Subsequently, 600 g / t of a compound collector was added, in which the ratio of black reagent to xanthate was 4:1. Finally, 20 g / t of pine oil was added, reacted for 2 minutes, and flotation was carried out for 5 minutes at natural pH and room temperature.
[0055] The tailings obtained from the first roughing process are then subjected to two more rounds of scavenging. The first round of scavenging involves adding 300 g / t of dispersant sodium hexametaphosphate and reacting for 5 minutes. Then, 300 g / t of compound collector is added and reacted for 5 minutes. Finally, 10 g / t of pine oil is added for a foaming reaction for 2 minutes, followed by flotation at natural pH and room temperature for 3 minutes. This round of scavenging yields the first batch of middlings and tailings. The middlings will be returned to the next roughing step, while the tailings will require further scavenging.
[0056] The second round of scavenging requires the addition of 150 g / t of dispersant sodium hexametaphosphate and a reaction time of 5 min; followed by the addition of 100 g / t of compound collector and a reaction time of 5 min; finally, 5 g / t of pine oil is added for a foaming reaction of 2 min, followed by flotation at natural pH and room temperature for 3 min; this round of scavenging yields a second batch of middlings and tailings, which will be returned to the first scavenging step of the next round. The results are shown in Table 2-2.
[0057] Table 4-1 Main elemental composition of mercury-removed roasted sand and sulfide gypsum residue
[0058] Table 4-2 Ag grade and recovery rate of mineral samples after agglomeration flotation
[0059] It is evident that in the absence of agglomerates, the quality of the concentrate and the two batches of middlings increased significantly, and the Ag grade decreased to 500-700 g / t. This indicates that a large amount of gangue entered the concentrate and middlings, causing the entire flotation process to fail.
[0060] The above embodiments are merely preferred embodiments of the present invention, used to illustrate the technical solutions of the present invention, and are not intended to limit the scope of protection of the present invention. Although the present invention has been described in detail with reference to preferred embodiments, various modifications, substitutions, or equivalent transformations can be made to its technical solutions without departing from the spirit and substance of the present invention. Any equivalent substitutions, improvements, or modifications made to the above embodiments based on the technical concept of the present invention should fall within the scope of protection of the present invention.
Claims
1. A method for recovering precious and base metals from smelter acid sludge, characterized in that, The method includes: (1) The smelting acid mud was oxidized and leached using a sulfuric acid-hydrogen peroxide leaching system to obtain a selenium-containing solution and gypsum residue; Compared to smelting acid sludge, the amount of hydrogen peroxide added is 1-5 wt%, and the amount of sulfuric acid added is 300-500 g / L; the liquid-to-solid ratio is 5-10:
1. (2) The selenium-containing solution was reduced with a reducing agent to obtain crude selenium; (3) The gypsum residue obtained in step (1) is pre-sulfurized, and then the solid and liquid are separated to obtain the sulfurized gypsum residue; (4) The silver and other precious metals in the sulfided gypsum slag obtained in step (3) are enriched by agglomeration flotation, including: primary roughing, primary scavenging, and secondary scavenging. The primary roughing yields silver concentrate and primary tailings. The primary tailings are subjected to primary scavenging to obtain primary intermediate ore and secondary tailings. The secondary tailings are subjected to secondary scavenging to obtain secondary intermediate ore and secondary tailings. The primary intermediate ore is returned to the next round of primary roughing, and the secondary intermediate ore is returned to the next round of primary scavenging. In the first roughing, first scavenging, and second scavenging processes, dispersants, agglomerants, collectors, and foaming agents are added sequentially, respectively.
2. The process for recovering and smelting the precious and noble metals from acid sludge according to claim 1, characterized in that, In step (1), before oxidative leaching, the following steps are also included: crushing and grinding the smelting acid mud to below 200 mesh.
3. The process for recovering and smelting the precious and non-precious metals from acid sludge as claimed in claim 1 wherein, In step (1), the leaching time is 2-6 hours.
4. The method for recovering rare and precious metals from smelting acid sludge according to claim 1, characterized in that, In step (2), the selenium-containing solution is heated and reduced with a reducing agent at 60-80℃; In step (2), the reducing agent is hydrazine hydrate and / or sodium sulfite.
5. The process for recovering and smelting the precious and non-precious metals from acid sludge as claimed in claim 1 wherein, In step (3), sodium sulfide solution is used for pre-sulfurization, the sulfidation time is 2-10 min, and the amount of sodium sulfide added is 500-1000 g / t gypsum residue.
6. The process for recovering and smelting the precious and non-precious metals from acid sludge as claimed in claim 1, wherein, In step (3), the slurry concentration of gypsum slag is 20-30%.
7. The process for recovering and smelting the precious and non-precious metals from acid sludge as claimed in claim 1, wherein, In step (4), at least one of the following characteristics must be satisfied: The dispersant is sodium hexametaphosphate; The agglomerating agent is emulsified kerosene; The composite collector is a mixture of black powder and yellow powder, with a mass ratio of 3-5:1; The foaming agent is pine oil.
8. The process for recovering and smelting the precious and non-precious metals from acid sludge as claimed in claim 1 wherein, Step (4) First coarse selection: The pulp concentration is 20-30%; The dosage of dispersant is 500-1000 g / t, and the reaction time is optionally 5-10 min; The amount of agglomerant added is 20-50 g / t, and the reaction time is optionally 15-30 min; The dosage of the compound collector is 300-600 g / t, and the reaction time is optionally 4-8 min; The dosage of foaming agent is 20-50 g / t, and the reaction time is optionally 1-3 min; The time for flotation and bubble scraping in a single roughing process at natural pH and room temperature can be arbitrarily set to 4-6 minutes.
9. The process for recovering and smelting the precious and non-precious metals from acid sludge as claimed in claim 1, wherein, Step (4) First scan selection: The dosage of dispersant is 200-600 g / t, and the reaction time is optionally 2-7 min; The amount of agglomerant added is 10-30 g / t, and the reaction time is optionally 10-20 min; The amount of compound collector added is 200-400 g / t, and the reaction time is optionally 3-6 min; The dosage of foaming agent is 10-20 g / t, and the reaction time is optionally 1-3 min; The time for the first scavenging flotation at natural pH and room temperature can be arbitrarily set to 2-4 minutes.
10. The process for recovering and smelting the precious and non-precious metals from acid sludge as claimed in claim 1 wherein, Step (4) Second scan selection: The dosage of dispersant is 100-300 g / t, and the reaction time is optionally 4-6 min; The amount of agglomerant added is 10-20 g / t, and the reaction time is optionally 10-20 min; The amount of compound collector added is 100-200 g / t, and the reaction time is optionally 4-6 min; The dosage of foaming agent is 5-10 g / t, and the reaction time is optionally 1-3 min; The second scavenging process, at natural pH and room temperature, can be performed for 2-4 minutes of flotation skimming.