Method for separating and recovering valuable metals from yellow slag

Abstract

The invention discloses a method for separating and recovering valuable metals from yellow slag, which comprises the following steps: preparing the yellow slag containing iron-arsenic-copper alloy into an anode, and performing electrochemical dissolution by using the anode to obtain anode slime and enriched iron electrolyte; sequentially sulfating, roasting and leaching the anode slime to obtain a leaching solution and leaching residues, and separating and recycling valuable metals in the leaching solution and the leaching residues through post-treatment; the mass of sulfuric acid used for sulfating is 35-60% of the mass of the anode mud. The invention provides a method for effectively separating valuable metals such as iron, copper, germanium and the like aiming at the problem that the yellow slag of the iron-arsenic-containing copper alloy is difficult to treat, the content of iron in the material can be effectively reduced through sulfating and roasting, the consumption of heavy acid in the subsequent treatment process is reduced, meanwhile, the risk of generating arsine by direct acid leaching is avoided, and the method is safer, more environmentally friendly and higher in economic benefit. And the separation effect of each valuable metal is good.

Description

Technical Field

[0001] This invention belongs to the field of resource recycling and reuse technology, and in particular relates to a method for separating and recovering valuable metals from yellow slag. Background Technology

[0002] Yellow slag is an arsenic-iron alloy formed during crude lead smelting, where iron and arsenic combine. Its melting point is higher than lead matte. It is trapped between copper matte slag and molten lead, forming a membrane that deteriorates furnace conditions. This leads to the formation of byproducts such as arsenic matte with copper and lead, while simultaneously dissolving rare and precious metals like gold, silver, and nickel. Therefore, regular cleaning is necessary. Due to differences in processes among different manufacturers, the specific composition and metal states of yellow slag vary considerably, making it difficult to process and separate and recover valuable metals using traditional methods. The main valuable metals, such as copper, germanium, nickel, and silver alloys, have significant economic value. For yellow slag, whose main component is an iron-arsenic-copper alloy and contains many rare and precious metals like germanium, indium, and silver, there are almost no existing processes for treating this material.

[0003] Some studies have employed high-temperature roasting and magnetic separation processes to treat arsenic matte, a material primarily composed of iron silicates or oxides. However, these processes, involving high temperatures, are insufficient for separating and recovering the slag, whose main component is an iron-arsenic-copper alloy. Furthermore, the equipment requirements are high, resulting in significant investment. Other studies have used oxygen pressure leaching to treat arsenic matte and recover valuable metals. However, this process requires high-pressure, high-temperature equipment, involves substantial investment, and the high copper content in the material, along with the potential for generating harmful gases such as arsine during acid leaching. Therefore, these methods offer low returns on investment for treating the aforementioned slag, which is primarily composed of an iron-arsenic-copper alloy. Summary of the Invention

[0004] In order to overcome at least one of the problems existing in the prior art, the object of the present invention is to provide a method for separating and recovering valuable metals from yellow slag.

[0005] To achieve the above objectives, the technical solution adopted by the present invention is as follows: This invention provides a method for separating and recovering valuable metals from yellow slag, comprising the following steps: making an anode from yellow slag containing an iron-arsenic-copper alloy; electrochemically dissolving the anode to obtain anode mud and iron-enriched electrolyte; sequentially sulfating, roasting, and leaching the anode mud to obtain a leachate and leaching residue; and separating and recovering the valuable metals therefrom through post-treatment; wherein the sulfuric acid used in the sulfation is 35-60% of the mass of the anode mud.

[0006] This invention targets yellow slag containing iron-arsenic-copper alloy (with iron-arsenic-copper alloy as the main component). It uses this alloy as an anode and employs electrochemical dissolution to treat the iron-arsenic-copper alloy, which is difficult to treat with conventional acid leaching or oxidative acid leaching. This reduces the need for large-scale crushing and grinding equipment, lowering investment and energy consumption. Simultaneously, it effectively reduces the iron content in the material, decreasing the consumption of heavy acid in subsequent processing and avoiding the risk of arsine formation from direct acid leaching. The remaining iron-arsenic-copper alloy material is enriched in the anode mud, and arsenic is separated through sulfation and roasting. Sulfation decomposes and destroys the alloy structure formed by iron, arsenic, and copper, forming sulfates. Roasting then causes the arsenic to form arsenic oxide and volatilize (sodium arsenate or calcium arsenate can be obtained through wet fixation). Simultaneously, high-temperature roasting decomposes the iron sulfate while the copper sulfate remains intact. This allows for the separation of iron from valuable metals such as copper and germanium during subsequent leaching, and reduces the problem of large amounts of iron ions in the solution being separated through precipitation during simple leaching, resulting in the loss of valuable metals such as copper and germanium. Sulfation using an appropriate amount of sulfuric acid can promote the formation of sulfates, especially the complete conversion of iron into iron sulfates, which are then decomposed into iron oxides through subsequent roasting, thus preventing iron from affecting the leaching of copper and germanium.

[0007] Preferably, the sulfuric acid used in the sulfation is 40-55% of the mass of the anode mud; more preferably, it is 45-55%.

[0008] Preferably, the total mass content of the iron-arsenic-copper alloy in the yellow slag is 87-99%; further, the yellow slag includes the following components by mass percentage: iron 66-74%, arsenic 16-24%, and copper 5-10%.

[0009] Preferably, the yellow residue also contains germanium; further, the germanium content is 0.1~0.5% by mass.

[0010] Preferably, the yellow residue also contains silver; further, the silver content is 0.01~0.05% by mass.

[0011] Preferably, the yellow slag comprises the following components by mass percentage: 66-74% iron, 16-24% arsenic, 5-10% copper, 0.1-0.5% germanium, and 0.01-0.1% silver.

[0012] Preferably, the electrolyte used in the electrochemical dissolution includes sulfate ions and iron ions; wherein the concentration of sulfate ions is 40~60 g / L and the concentration of iron ions is 15~35 g / L; more preferably, the concentration of sulfate ions is 45~55 g / L and the concentration of iron ions is 20~30 g / L.

[0013] Preferably, the voltage used for the electrochemical dissolution is 0.5~1.5V; more preferably 0.8~1.2V.

[0014] Preferably, 75-85% of the total mass of iron is dissolved in the electrolyte or enriched in the cathode by the electrochemical dissolution to obtain metallic iron; more preferably, 78-82%.

[0015] Preferably, the calcination holding temperature is 250~850℃; for example, it can be any value or a range between 250℃, 300℃, 400℃, 500℃, 600℃, 700℃, 800℃ or 850℃.

[0016] Preferably, the calcination holding time is 0.5 to 5 hours; for example, it can be any value of 0.5 hours, 1 hour, 2 hours, 3 hours, 4 hours or 5 hours or a range between any two.

[0017] Preferably, the heating rate of the calcination is 5~15℃ / min; more preferably 8~12℃ / min.

[0018] Preferably, the roasting includes a first roasting, a second roasting, and a third roasting performed sequentially; the first roasting is held at a temperature of 200~300℃ for 0.3~0.7h; the second roasting is held at a temperature of 450~650℃ for 0.3~1.5h; and the third roasting is held at a temperature of 650~850℃ for 0.3~1.5h.

[0019] Preferably, the holding temperature for the second roasting is 550~650℃.

[0020] Preferably, the holding temperature for the third roasting is 650~750℃.

[0021] By employing appropriate roasting processes, higher leaching rates of copper and germanium can be obtained, enabling the effective separation of iron from high-value metals such as copper and germanium.

[0022] Preferably, the leaching agent used in the leaching is water or an aqueous solution of sulfuric acid; more preferably, the leaching agent used in the leaching is water.

[0023] Preferably, the concentration of the sulfuric acid aqueous solution is 1~55 g / L; more preferably 25~50 g / L.

[0024] Preferably, the liquid-to-solid ratio used in the leaching is (3~7):1; more preferably (4~6):1.

[0025] Preferably, the leaching temperature is 80~100℃; more preferably 85~95℃.

[0026] Preferably, the leaching time is 0.5 to 1.5 hours; more preferably, it is 0.8 to 1.2 hours.

[0027] Preferably, the leachate contains germanium and copper; the post-processing includes the following steps: mixing the leachate with tartaric acid for reaction, and then adding a germanium extractant for extraction and separation to obtain a germanium-containing solution and a copper-containing solution.

[0028] Preferably, the mass of the tartaric acid is 3 to 5 times the germanium content in the leachate; more preferably, it is 3.5 to 4.5 times.

[0029] Preferably, the germanium extractant includes at least one of LIX63, N235, or TBP.

[0030] Preferably, the pH of the copper-containing solution is adjusted to 6-7 using a pH adjuster to obtain basic copper carbonate; more preferably, the pH of the copper-containing solution is adjusted to 6.3-6.7 using a pH adjuster.

[0031] Preferably, the pH adjuster includes sodium hydroxide, sodium carbonate, or a combination thereof.

[0032] Preferably, the post-processing includes the following steps: smelting the leaching residue with lead and zinc to recover the remaining valuable metals.

[0033] Preferably, the enriched iron electrolyte is concentrated and crystallized to obtain an iron-containing salt; further, the iron-containing salt includes ferrous sulfate.

[0034] The beneficial effects of this invention are as follows: This invention addresses the problem of difficult treatment of yellow slag containing iron, arsenic, and copper alloys by providing an effective method for separating valuable metals such as iron, copper, and germanium. Through sulfation and roasting, the iron content in the material can be effectively reduced, reducing the consumption of heavy acid in subsequent processing. At the same time, it avoids the risk of arsine formation from direct acid leaching, making it safer and more environmentally friendly, and the separation effect of each valuable metal is good. Attached Figure Description

[0035] Figure 1 This is a flowchart of the method for separating and recovering valuable metals from yellow slag in Embodiment 1 of the present invention.

[0036] Figure 2 This is a photograph of the yellow residue and ferrous sulfate crystals involved in Example 1 of the present invention. Detailed Implementation

[0037] The following specific embodiments further illustrate the content of the present invention in detail. It should also be understood that the following embodiments are only for further explanation of the present invention and should not be construed as limiting the scope of protection of the present invention. Non-essential improvements and adjustments made by those skilled in the art based on the principles described herein are all within the scope of protection of the present invention. The specific process parameters, etc., in the following examples are merely examples within a suitable range; that is, those skilled in the art can make selections within a suitable range based on the description herein, and are not intended to be limited to the specific data in the examples below. Unless otherwise specified, the raw materials, reagents, or apparatus used in the following embodiments and comparative examples can be obtained from conventional commercial sources or by existing known methods.

[0038] In the following examples and comparative examples, the composition of the yellow slag used, by mass percentage, is as follows: iron 70%, arsenic 20%, copper 9.8%, germanium 0.19%, and silver 0.01%.

[0039] Example 1 A method for separating and recovering valuable metals from yellow slag, the flowchart of which is shown below. Figure 1 As shown, the specific steps are as follows: Yellow slag is cast into an anode plate, which is then placed in a titanium basket for electrochemical dissolution (electrolysis). The cathode is made of stainless steel or iron, and the electrolyte is a ferric sulfate solution with a sulfate ion concentration of 50 g / L and an iron ion concentration of 25 g / L, at a voltage of 1.0 V. Through electrochemical dissolution, 80% of the iron by mass is introduced into the electrolyte or enriched at the cathode to obtain metallic iron. After electrolysis, the iron-enriched electrolyte is concentrated and crystallized to obtain ferrous sulfate crystals. The concentrated solution is then leached back into the lead-zinc smelting system for enrichment. Collect any remaining rare and dispersed metals such as germanium. Collect the anode mud after electrolysis. Take 50g of anode mud and add concentrated sulfuric acid (50% by weight of the anode mud) and stir. After stirring evenly, place it in a muffle furnace for calcination. Increase the temperature at 10℃ / min, from room temperature to 250℃ and hold for 0.5h, then increase the temperature to 600℃ and hold for 1h, then increase the temperature to 700℃ and hold for 1h. Add 250mL of pure water for leaching, with a liquid-to-solid ratio of 5:1, at 90℃, stirring at 200rpm, for leaching for 1h. After leaching, filter to obtain the leachate and leaching residue. The leachate is rich in copper and germanium. Tartaric acid (additive) with a germanium content of 4 times is added to the leachate and reacted for 0.5 hours. Germanium is separated by N235 extraction with a germanium extraction rate of 92%. The raffinate is a copper-containing solution. Its pH is adjusted to 6.5 with Na2CO3 to prepare basic copper carbonate. The leaching residue is rich in iron (such as iron oxide) and is returned to the lead-zinc smelting system to enrich and recover the residual metal.

[0040] The content of valuable metals in the leachate was analyzed by sampling. The results showed that the leaching rate of iron was 2.16%, the leaching rate of copper was 92.6%, and the leaching rate of germanium was 85.8%.

[0041] Example 2 A method for separating and recovering valuable metals from yellow slag differs from Example 1 in that the roasting process is modified as follows: the temperature is increased from room temperature to 250°C and held for 0.5 h at a heating rate of 10°C / min, then increased to 600°C and held for 1 h, and then increased to 800°C and held for 0.5 h; other conditions are the same as in Example 1.

[0042] The contents of valuable metals were analyzed by sampling the leachate. The results showed that the leaching rate of iron was 0.26%, the leaching rate of copper was 85.2%, and the leaching rate of germanium was 86.2%.

[0043] Example 3 A method for separating and recovering valuable metals from yellow slag differs from Example 1 in that concentrated sulfuric acid, accounting for 40% of the mass of the anode mud, is added and stirred; the roasting process is modified as follows: the temperature is increased from room temperature to 250°C at a rate of 10°C / min and held for 0.5 hours, then increased to 500°C and held for 0.5 hours, and then increased to 700°C and held for 1 hour; other conditions are the same as in Example 1.

[0044] The content of valuable metals in the leachate was analyzed by sampling. The results showed that the leaching rate of iron was 5.5%, the leaching rate of copper was 85.6%, and the leaching rate of germanium was 35.8%.

[0045] Comparative Example 1 A method for separating and recovering valuable metals from yellow slag differs from Example 1 in that concentrated sulfuric acid, accounting for 30% of the mass of the anode mud, is added and stirred; the roasting process is modified to: heating from room temperature to 250°C at a rate of 10°C / min and holding for 0.5 hours, then heating to 500°C and holding for 0.5 hours, followed by heating to 700°C and holding for 0.5 hours; the leaching agent used is modified to a 25 g / L sulfuric acid aqueous solution; other conditions are the same as in Example 1.

[0046] The content of valuable metals in the leachate was analyzed by sampling. The results showed that the leaching rate of iron was 28.1%, the leaching rate of copper was 0.3%, and the leaching rate of germanium was 0.6%.

[0047] Comparative Example 2 A method for separating and recovering valuable metals from yellow slag differs from Example 1 in that concentrated sulfuric acid, accounting for 20% of the mass of the anode mud, is added and stirred; the roasting process is modified to: heating from room temperature to 500℃ at a rate of 10℃ / min and holding for 0.5h, then heating to 600℃ and holding for 2h; the leaching agent used is modified to a 50g / L sulfuric acid aqueous solution; other conditions are the same as in Example 1.

[0048] The content of valuable metals in the leachate was analyzed by sampling. The results showed that the leaching rate of iron was 57.3%, the leaching rate of copper was 0.06%, and the leaching rate of germanium was 0.08%.

[0049] Comparative Example 3 A method for separating and recovering valuable metals from yellow slag, comprising the following steps: Take 50g of yellow residue, grind it to 120 mesh, leach it with 2mol / L sulfuric acid, add 10% hydrogen peroxide, temperature 90~95℃, liquid-solid ratio 5:1, leaching time 4h, filter after leaching to obtain leachate and leach residue.

[0050] Analysis of samples taken from the leachate revealed the following: iron leaching rate was 29.3%, with most iron and copper remaining unleached; the leaching residue weighed 42.6g, and arsine was generated during the leaching process.

[0051] Comparative Example 4 A method for separating and recovering valuable metals from yellow slag, comprising the following steps: Take 50g of yellow residue and grind it to 120 mesh. Leach it with 5mol / L hydrochloric acid, add 10% hydrogen peroxide, maintain a temperature of 90~95℃, a liquid-to-solid ratio of 5:1, and a leaching time of 4h. After leaching, filter to obtain leachate and leach residue.

[0052] Analysis of samples taken from the leachate revealed the following: iron leaching rate was 23%, with most iron and copper remaining unleached; the leaching residue weighed 44.0g.

[0053] The leaching rates of the above embodiments and comparative examples are summarized in Table 1 below.

[0054] Table 1. Comparison of leaching rates between various embodiments and comparative examples.

[0055] As shown in Table 1, the methods of Examples 1-3 of this invention can effectively separate and recover valuable metals such as iron, copper, and germanium. The leaching rates of copper and germanium in Examples 1-3 are relatively high, while iron is recovered by decomposing into oxides during high-temperature roasting, resulting in a lower leaching rate in the leaching step, which is more conducive to the recovery and separation of copper and germanium. In Comparative Examples 1-2, the amount of sulfuric acid added during the sulfation of the anode mud was insufficient. Unreacted iron reduced the leached copper and germanium, leading to a high iron leaching rate and a low copper and germanium leaching rate, affecting subsequent separation. This resulted in a large amount of iron ions in the solution being separated by precipitation, causing the entrainment loss of valuable metals such as copper and germanium. Comparative Examples 3-4 used conventional oxidative acid leaching processes, which were difficult to leach most of the iron and copper, and arsine was generated during the leaching process.

[0056] Figure 2The images shown are actual photos of yellow slag and ferrous sulfate crystals involved in Example 1 of this invention; where A represents yellow slag and B represents ferrous sulfate crystals. It can be seen that the method of this invention can separate and recover valuable metals from yellow slag, thereby obtaining high-purity recovered products, such as ferrous sulfate crystals.

[0057] This invention targets yellow slag containing iron-arsenic-copper alloy (with iron-arsenic-copper alloy as the main component). It uses this alloy as an anode and employs electrochemical dissolution to treat the iron-arsenic-copper alloy, which is difficult to treat with conventional acid leaching or oxidative acid leaching. This reduces the need for large-scale crushing and grinding equipment, lowering investment and energy consumption. Simultaneously, it effectively reduces the iron content in the material, decreasing the consumption of heavy acid in subsequent processing and avoiding the risk of arsine formation from direct acid leaching. The remaining iron-arsenic-copper alloy material is enriched in the anode mud, and arsenic is separated through sulfation and roasting. Sulfation decomposes and destroys the alloy structure formed by iron, arsenic, and copper, forming sulfates. Roasting then causes the arsenic to form arsenic oxide and volatilize (sodium arsenate or calcium arsenate can be obtained through wet fixation). Simultaneously, high-temperature roasting decomposes the iron sulfate while the copper sulfate remains intact. This allows for the separation of iron from valuable metals such as copper and germanium during subsequent leaching, and reduces the problem of large amounts of iron ions in the solution being separated through precipitation during simple leaching, resulting in the loss of valuable metals such as copper and germanium.

[0058] In summary, this invention addresses the problem of difficult treatment of yellow slag containing iron, arsenic, and copper alloys by providing an effective method for separating valuable metals such as iron, copper, and germanium. Through sulfation and roasting, the iron content in the material can be effectively reduced, thus reducing the consumption of heavy acid in subsequent processing. At the same time, it avoids the risk of arsine formation from direct acid leaching, making it safer and more environmentally friendly, and achieving good separation of each valuable metal.

Claims

1. A method for separating and recovering valuable metals from yellow slag, characterized in that, Includes the following steps: Yellow slag containing iron, arsenic and copper alloy was made into an anode, and the anode was used for electrochemical dissolution to obtain anode mud and iron-enriched electrolyte. The anode mud is subjected to sulfation, roasting, and leaching in sequence to obtain leachate and leaching residue. Valuable metals are then recovered through post-treatment. The sulfuric acid used in the sulfation is 35-60% of the mass of the anode mud.

2. The method according to claim 1, characterized in that, The yellow slag comprises the following components by mass percentage: 66-74% iron, 16-24% arsenic, 5-10% copper, 0.1-0.5% germanium, and 0.01-0.1% silver.

3. The method according to claim 1, characterized in that, The electrolyte used in the electrochemical dissolution includes sulfate ions and iron ions; wherein the concentration of sulfate ions is 40~60 g / L and the concentration of iron ions is 15~35 g / L. And / or, the voltage used for the electrochemical dissolution is 0.5~1.5V; And / or, by means of electrochemical dissolution, 75-85% of the total mass of iron is dissolved in the electrolyte or enriched in the cathode to obtain metallic iron.

4. The method according to claim 1, characterized in that, The roasting temperature is 250~850℃; And / or, the calcination holding time is 0.5~5h; And / or, the heating rate of the calcination is 5~15℃ / min.

5. The method according to claim 1, characterized in that, The roasting process includes a first roasting, a second roasting, and a third roasting performed sequentially; the first roasting is held at a temperature of 200~300℃ for 0.3~0.7h; the second roasting is held at a temperature of 450~650℃ for 0.3~1.5h; and the third roasting is held at a temperature of 650~850℃ for 0.3~1.5h.

6. The method according to claim 1, characterized in that, The leaching agent used is water or an aqueous sulfuric acid solution; the concentration of the aqueous sulfuric acid solution is 1~55 g / L; And / or, the liquid-to-solid ratio used in the leaching is (3~7):1; And / or, the leaching temperature is 80~100℃; And / or, the leaching time is 0.5 to 1.5 hours.

7. The method according to claim 1, characterized in that, The leachate contains germanium and copper; the post-treatment includes the following steps: The leachate was mixed with tartaric acid and reacted, and then germanium extractant was added for extraction and separation to obtain germanium-containing solution and copper-containing solution.

8. The method according to claim 7, characterized in that, The mass of the tartaric acid is 3 to 5 times the germanium content in the leachate; And / or, the germanium extractant includes at least one of LIX63, N235 or TBP; And / or, the pH of the copper-containing solution is adjusted to 6-7 using a pH adjuster to obtain basic copper carbonate; the pH adjuster includes sodium hydroxide, sodium carbonate, or a combination thereof.

9. The method according to claim 1, characterized in that, The post-processing includes the following steps: smelting the leaching residue with lead and zinc to recover the remaining valuable metals.

10. The method according to claim 1, characterized in that, The enriched iron electrolyte was concentrated and crystallized to obtain an iron-containing salt.

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