A method for separating and enriching thallium in zinc oxide ash

Abstract

The present invention belongs to the field of recovery of zinc oxide ash from wet zinc smelting, and discloses a method for separating and enriching thallium from zinc oxide ash. The method comprises the following steps: (1) grinding and mixing zinc oxide ash and sodium hydroxide or potassium hydroxide in a mass ratio of 3 to 5:1, and calcining at a temperature of 200 to 350°C to obtain a calcined product; (2) slurrying the calcined product obtained in step (1) with water in a solid-liquid ratio of 1:3 to 1:6, and then leaching the slurry in water at a temperature of 60 to 80°C, with an end point pH value of 11 to 12, filtering, and obtaining a water leaching liquid and a water leaching residue; (3) subjecting the water leaching residue obtained in step (2) to low-acid leaching at a temperature of 70 to 90°C, with an end point pH value of 3.5 to 4.5, filtering, and obtaining an acid leaching liquid and an acid leaching residue enriched with thallium. The method has a simple process, low cost, and does not disperse thallium. The method has good separation and enrichment effects, thereby achieving comprehensive resource utilization, avoiding internal circulation of thallium, and reducing the thallium removal pressure of the zinc system.

Description

Technical Field

[0001] The invention belongs to the field of zinc oxide ash recovery in hydrometallurgy zinc smelting, and particularly relates to a method for separating and enriching thallium in zinc oxide ash. Background Art

[0002] Zinc oxide soot from hydrometallurgical zinc smelting is mostly produced during the defluorination and chlorination process in a multi-hearth furnace. Substances with lower volatilization temperatures are concentrated in the zinc oxide soot, which primarily contains elements such as zinc, lead, cadmium, thallium, fluorine, and chlorine. The difficulty in recycling zinc oxide soot lies in the maximum separation of fluorine, chlorine, and valuable metals such as zinc, as well as the separation and recovery of the highly toxic element thallium. Due to the complexity and high cost of zinc oxide soot recovery, most zinc smelters currently export or stockpile their zinc oxide soot, with no secondary use, let alone thallium disposal. For example, a hydrometallurgical zinc smelting plant with an annual output of 300,000 tons produces 800 to 900 tons of zinc oxide soot annually. Exporting or stockpiling zinc oxide carries significant safety and environmental pressures, and there is a risk of leakage.

[0003] The current conventional treatment process for zinc oxide fly ash involves alkali washing to separate fluorine and chlorine, followed by acid leaching of the alkali wash residue to recover metals such as zinc and cadmium. Thallium is removed from the alkali wash solution by sulfidation precipitation, and thallium is concentrated in the sulfided slag, which is generally stored as a relatively stable material. However, the fluorine and chlorine removal rate during the alkali wash process is generally low, at only 50-60%. The alkali wash residue cannot meet the requirements for secondary use, and the thallium recovery rate during the alkali wash process is limited. Some thallium will be dispersed in the alkali wash residue, which circulates through the hydrometallurgical zinc smelting system as it leaches out, putting pressure on the system's stable operation. Furthermore, a corresponding thallium removal process must be used to reduce the thallium content of the zinc sulfate solution, increasing operating costs. Summary of the Invention

[0004] In view of the above problems existing in the prior art, the object of the present invention is to provide a method for separating and enriching thallium in zinc oxide ash, which has simple process, low cost, does not disperse thallium, and has good separation and enrichment effect.

[0005] To achieve the above object, the technical solution adopted by the present invention is:

[0006] A method for separating and enriching thallium in zinc oxide ash comprises the following steps:

[0007] (1) Grind and mix zinc oxide soot and sodium hydroxide or potassium hydroxide in a mass ratio of 3 to 5:1, and calcine at a temperature of 200 to 350°C to obtain a calcined product;

[0008] (2) slurrying the roasted product obtained in step (1) with water at a solid-liquid ratio of 1:3 to 1:6, and then soaking in water at a temperature of 60 to 80°C and an end point pH of 11 to 12, filtering to obtain a water soak and a water soak residue;

[0009] (3) The water leaching residue obtained in step (2) is subjected to low-acid leaching at a temperature of 70-90° C. and an end-point pH value of 3.5-4.5, and filtered to obtain an acid leaching solution and an acid leaching residue enriched with thallium.

[0010] The present invention creatively adopts sodium alkali roasting to solidify thallium in zinc oxide soot, so that thallium is not dispersed during the separation and enrichment process.

[0011] The reaction principle involved is that the physical phases of thallium in zinc oxide soot are thallium oxide and thallium chloride. The curves of the vapor pressure of thallium oxide and thallium chloride changing with temperature are as follows: Figure 1 shown.

[0012] When the temperature is below 350°C, the volatility of the two thallium compounds is very low and they remain in the burned product. When the sodium alkali of the present invention is used for roasting, the following chemical reaction occurs:

[0013] TlCl+3NaOH=TlOH+NaCl (1)

[0014] 2TlOH=Tl2O+H2O (2)

[0015] Tl2O+O2=Tl2O3(3)

[0016] During water leaching, the pH value of the present invention is controlled at 11-12, so that thallium (thallium trioxide), zinc, lead, cadmium and the like will not be leached, while fluorine and chlorine are leached into the solution to the maximum extent, thereby realizing the first step of separation and enrichment of thallium; during acid leaching, the pH value is controlled at 3.5-4.5, so that thallium is hydrolyzed in the form of thallium hydroxide and coexists with lead sulfate in the acid leaching residue, while metals such as zinc and cadmium are leached into the solution, thereby realizing the further separation and enrichment of thallium.

[0017] Preferably, in step (1), the heating rate of the calcination is 5-10°C / min; the holding time is 60-120 min; and the mixture is naturally cooled to room temperature after the holding is completed.

[0018] Preferably, in step (2), the sintered blocks in the calcined product are crushed and ground before slurrying; the particle size of the calcined product after crushing and grinding is -200 mesh accounting for ≥80%.

[0019] Preferably, in step (2), the immersion time is 1 to 3 hours.

[0020] Preferably, in step (2), the water extract is evaporated and crystallized to recover sodium chloride / potassium and sodium fluoride / potassium.

[0021] Preferably, in step (3), the water-leached residue is mixed with sulfuric acid or electrolytic waste liquid to perform low-acid leaching.

[0022] Preferably, in step (3), the solid-liquid ratio of low-acid leaching is 1:3-1:5; and the initial acid is 60-80 g / l.

[0023] Preferably, in step (3), the low-acid leaching time is 1 to 3 hours.

[0024] Preferably, in step (3), the pickling solution is sent to a hydrometallurgical zinc smelting system.

[0025] Compared with the prior art, the present invention has the following beneficial effects:

[0026] (1) The present invention adopts a low-temperature sodium alkali roasting method to fix the thallium in the zinc oxide ash, so that it forms relatively stable thallium trioxide that is not easily leached out by alkali, thereby preventing the thallium from being dispersed during the water washing defluorination and dechlorination process.

[0027] (2) The present invention increases the activity of fluorine and chlorine through low-temperature roasting, and improves their removal rate in the water washing process. The water-washed slag can contain 0.013% fluorine and 0.008% chlorine. The fluorine removal rate is greater than 95% and the chlorine removal rate is greater than 99%, meeting the requirements of secondary utilization.

[0028] (3) The present invention adopts low-temperature sodium alkali roasting, water leaching and low-acid leaching to remove fluorine, chlorine, zinc and cadmium respectively, so that thallium is enriched, which greatly reduces the amount of thallium-containing materials, facilitates storage and transportation, and the next step of thallium recovery and disposal.

[0029] (4) The process of the present invention is simple and low in cost. Thallium is not dispersed and has good separation and enrichment effects. It not only realizes the comprehensive utilization of resources, but also avoids the internal circulation of thallium, and reduces the thallium removal pressure of the zinc system. BRIEF DESCRIPTION OF THE DRAWINGS

[0030] The accompanying drawings are used to provide a further understanding of the present invention and constitute a part of the specification. Together with the embodiments of the present invention, they are used to explain the present invention and do not constitute a limitation of the present invention. In the accompanying drawings:

[0031] Figure 1 is a graph showing the change in vapor pressure of thallium oxide and thallium chloride with temperature;

[0032] Figure 2 It is a process flow chart of the present invention. DETAILED DESCRIPTION

[0033] To facilitate understanding of the present invention, the present invention will be described in more comprehensive and detailed form below in conjunction with the accompanying drawings and preferred embodiments. However, the protection scope of the present invention is not limited to the following specific embodiments.

[0034] Unless otherwise defined, all technical terms used hereinafter have the same meanings as those generally understood by those skilled in the art. The technical terms used herein are only for the purpose of describing specific embodiments and are not intended to limit the scope of protection of the present invention.

[0035] Unless otherwise specified, various raw materials, reagents, instruments and equipment used in the present invention can be purchased from the market or prepared by existing methods.

[0036] Example 1

[0037] This embodiment discloses a method for separating and enriching thallium in zinc oxide soot, such as Figure 2 As shown, the following steps are included:

[0038] (1) Sodium alkali roasting: Grind and mix zinc oxide soot and sodium hydroxide in a mass ratio of 3:1, spread the mixture in a crucible, and roast at 200°C for 120 min to obtain the roasted product;

[0039] (2) Water immersion: The sintered blocks in the roasted product were crushed and ground in a mortar. The crushed roasted product was slurried with water at a solid-liquid ratio of 1:3 and then water-immersed at 80°C for 3 h. The end point pH was 11. The product was filtered and the water immersion liquid was evaporated and crystallized to recover sodium chloride and sodium fluoride. The fluorine removal rate was calculated to be 95.05% and the chlorine removal rate was 99.14%.

[0040] (3) Low-acid leaching: Mix the water leaching residue with dilute sulfuric acid, control the initial acid to 65 g / L, and the liquid-solid ratio to 3:1 for low-acid leaching. The temperature is 70 ° C, the time is 3 h, and the end point pH is 4.5. Filter, zinc and other valuable metals enter the acid leaching solution. The calculated zinc leaching rate is 98.52%, and thallium is enriched in the acid leaching residue.

[0041] Example 2

[0042] This embodiment discloses a method for separating and enriching thallium in zinc oxide soot, such as Figure 2 As shown, the following steps are included:

[0043] (1) Sodium alkali roasting: Grind and mix zinc oxide soot and potassium hydroxide in a mass ratio of 5:1, spread the mixture in a crucible, and roast at 300 °C for 90 min to obtain the roasted product;

[0044] (2) Water immersion: The sintered blocks in the roasted product were crushed and ground in a mortar. The crushed roasted product was slurried with water at a solid-liquid ratio of 1:4 and then immersed in water at a temperature of 75°C for 2 h. The end point pH was 12. The product was filtered and the water immersion liquid was evaporated and crystallized to recover sodium chloride and sodium fluoride. The fluorine removal rate was calculated to be 95.89% and the chlorine removal rate was 99.51%.

[0045] (3) Low acid leaching: The water leaching residue was mixed with dilute sulfuric acid, the initial acid was controlled to 70 g / L, and the liquid-solid ratio was 4:1 for low acid leaching. The temperature was 75 ° C, the time was 3 h, and the end point pH was 4.0. After filtration, zinc and other valuable metals entered the acid leaching solution. The calculated zinc leaching rate was 98.73%, and thallium was enriched in the acid leaching residue.

[0046] Example 3

[0047] This embodiment discloses a method for separating and enriching thallium in zinc oxide soot, such as Figure 2 As shown, the following steps are included:

[0048] (1) Sodium alkali roasting: Grind and mix zinc oxide soot and sodium hydroxide in a mass ratio of 4:1, spread the mixture in a crucible, and roast at 350 ° C for 70 min to obtain the roasted product;

[0049] (2) Water immersion: The sintered blocks in the roasted product were crushed and ground in a mortar. The crushed roasted product was slurried with water at a solid-liquid ratio of 1:5 and then immersed in water at a temperature of 70°C for 3 hours. The end point pH was 12. The product was filtered and the water immersion liquid was evaporated and crystallized to recover sodium chloride and sodium fluoride. The fluorine removal rate was calculated to be 95.37% and the chlorine removal rate was 99.2%.

[0050] (3) Low acid leaching: The water leaching residue was mixed with the electrolytic waste liquid, the initial acid was controlled to 80 g / L, and the liquid-solid ratio was 5:1 for low acid leaching. The temperature was 90 ° C, the time was 3 h, and the end point pH was 3.5. After filtration, zinc and other valuable metals entered the acid leaching solution. The calculated zinc leaching rate was 98.90%, and thallium was enriched in the acid leaching residue.

[0051] Comparative Example 1

[0052] Compared with Example 1, in step (1), the sodium alkali roasting temperature is 400° C., and other method steps are the same.

[0053] Comparative Example 2

[0054] Compared with Example 2, in step (2), the water immersion temperature is 40° C., and other method steps are the same.

[0055] Comparative Example 3

[0056] Compared with Example 3, in step (3), the end point pH of the acid leaching is 2.5. Other than that, the method steps are the same.

[0057] The samples obtained in Examples 1-3 and Comparative Examples 1-3 were tested for fluorine, chlorine, zinc, and thallium components to calculate the fluorine and chlorine removal rates, zinc leaching rate, and thallium retention rate. The results are shown in Table 1.

[0058] Table 1

[0059] name Fluorine removal rate (%) Chlorine removal rate (%) Zinc leaching rate (%) Thallium retention rate (%) Example 1 95.05 99.14 98.52 99.58 Example 2 95.89 99.51 98.73 99.41 Example 3 95.37 99.2 98.90 99.35 Comparative Example 1 95.71 99.32 98.46 95.29 Comparative Example 2 90.15 94.82 98.57 97.64 Comparative Example 3 95.33 99.17 98.65 82.29

[0060] As can be seen from the above table, the method adopted in the embodiment of the present invention can separate thallium from fluorine, chlorine and zinc, and thallium is enriched in the acid leaching residue. Various technical indicators are significantly better than those of the comparative example.

[0061] The above are merely preferred embodiments of the present invention and are not intended to limit the present invention. Those skilled in the art will readily appreciate that the present invention is susceptible to various modifications and variations. Any modifications, equivalent substitutions, or improvements made within the spirit and principles of the present invention are intended to be within the scope of the present invention.

Claims

1. A method for separating and enriching thallium in zinc oxide ash, characterized in that: The following steps are involved: (1) Grind and mix zinc oxide soot with sodium hydroxide or potassium hydroxide in a mass ratio of 3 to 5:1, and calcine at a temperature of 200 to 350°C to obtain a calcined product; (2) slurrying the roasted product obtained in step (1) with water at a solid-liquid ratio of 1:3 to 1:6, and then soaking in water at a temperature of 60 to 80°C and an end point pH of 11 to 12, filtering to obtain a water soak and a water soak residue; (3) The water leaching residue obtained in step (2) is subjected to low-acid leaching at a temperature of 70-90° C. and an end-point pH value of 3.5-4.5, and filtered to obtain an acid leaching solution and an acid leaching residue enriched with thallium.

2. The method according to claim 1, wherein In step (1), the heating rate of the roasting is 5-10°C / min; the holding time is 60-120min; and the mixture is naturally cooled to room temperature after the holding is completed.

3. The method according to claim 1, wherein In step (2), the sintered blocks in the calcined product are crushed and ground before slurrying; the particle size of the calcined product after crushing and grinding is -200 mesh, accounting for ≥80%.

4. The method according to claim 1 or 2, wherein: In step (2), the immersion time is 1 to 3 hours.

5. The method according to claim 1, wherein In step (2), the water extract is evaporated and crystallized to recover sodium chloride / potassium and sodium fluoride / potassium.

6. The method according to claim 1, wherein In step (3), the water-leached residue is mixed with sulfuric acid or electrolytic waste liquid to perform low-acid leaching.

7. The method according to claim 1 or 6, wherein: In step (3), the solid-liquid ratio of low-acid leaching is 1:3~1:5; the initial acid is 60~80g / l.

8. The method according to claim 1, wherein In step (3), the low-acid leaching time is 1 to 3 hours.

9. The method according to claim 1, wherein In step (3), the pickling solution is sent to the hydrometallurgical zinc smelting system.

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