Process for recovering tin and uranium by wet process-pyrogenic process combined treatment of radioactive tantalum-niobium-containing slag

Abstract

The invention discloses a process for recovering tin and uranium by wet process-pyrogenic process combined treatment of radioactive tantalum-niobium slag, which comprises the following steps: (1) taking the tantalum-niobium slag containing tin and uranium as a raw material, adding an acid leaching agent into the raw material for leaching treatment, enriching tin in a solid phase to form tin-enriched slag, dissolving uranium in a liquid phase to form uranium-containing leachate, and leaching the uranium-enriched slag to obtain uranium-enriched slag; then carrying out solid-liquid separation to obtain tin-enriched slag and uranium-containing leachate; and (2) a cosolvent and a reducing agent are added into the tin-enriched slag obtained in the step (1), heat-preservation smelting is carried out, tin is reduced and separated from a slag phase, and a crude tin product and smelting slag are obtained after cooling. According to the method, through the two-step combined process of low-acid wet method tin leaching and enrichment and high-temperature pyrogenic process crude tin extraction, effective utilization of tin, uranium and other resources in the tantalum-niobium slag can be achieved, the recovery rate is high, the process is short and simple, the cost is low, and meanwhile reduction and recycling of the tantalum-niobium tailings are achieved.

Description

Technical Field

[0001] This invention belongs to the field of hazardous waste resource utilization and valuable metal recovery technology, and particularly relates to a process for treating radioactive tantalum and niobium slag and recovering tin and uranium. Background Technology

[0002] The smelting of tantalum and niobium generates a large amount of tantalum and niobium waste residue. Leaching with a mixture of sulfuric acid and hydrofluoric acid can effectively decompose tantalum and niobium concentrate. The leaching residue generated in this process is classified as hazardous waste, with its main elemental composition being Fe, S, Si, O, and F, along with a small amount of radioactive U. Direct stockpiling or disposal of this residue can easily cause environmental pollution and radioactive risks. At the same time, the waste residue contains a certain amount of valuable metal Sn. From the perspective of tin smelting raw materials, there is no raw material purchase cost, which is a significant advantage over ordinary tin smelters. It is a secondary resource with recycling value. Furthermore, the recovery of uranium can significantly reduce the amount of hazardous waste residue and reduce waste disposal or stockpiling costs.

[0003] Current processes for treating tantalum-niobium slag are relatively simple, mainly because the presence of uranium poses a risk to the recovery of elements such as tin, tantalum, and niobium, thus neglecting the extraction of elements like tin. Therefore, there is an urgent need to develop a process that achieves high recovery rates, low costs, and can simultaneously recover elements such as uranium and tin, overcoming the shortcomings of existing technologies. Summary of the Invention

[0004] The technical problem to be solved by the present invention is to overcome the deficiencies and defects mentioned in the background art above, and to provide a process for recovering tin and uranium from radioactive tantalum-niobium slag by a combination of wet and pyrometallurgical processes, which has high recovery rate and low cost.

[0005] To solve the above-mentioned technical problems, the technical solution proposed by this invention is as follows: A combined wet-pyrometallurgical process for recovering tin and uranium from radioactive tantalum-niobium slag includes the following steps: (1) Using tin- and uranium-containing tantalum-niobium slag as raw material, acid leaching agent is added to the raw material for leaching treatment, so that tin is enriched in the solid phase to form tin-enriched slag, and uranium is dissolved in the liquid phase to form uranium-containing leachate. Then, solid-liquid separation is performed to obtain tin-enriched slag and uranium-containing leachate. The acid leaching agent is at least one of sulfuric acid, hydrochloric acid, and organic acid, and the concentration of the acid leaching agent is 0.3-2 mol / L; (2) Add flux and reducing agent to the tin enrichment slag obtained in step (1), and heat-hold smelting at a temperature of 1200-1500℃ to reduce and separate tin from the slag phase. After cooling, crude tin product and smelting slag are obtained. The co-solvent is selected from at least one of CaF2, NaF, and Na2CO3; the reducing agent is elemental carbon.

[0006] As a further improvement, the tin- and uranium-containing tantalum-niobium slag in step (1) is tailings obtained by smelting tantalum-niobium ore using the hydrofluoric acid-sulfuric acid method.

[0007] As a further improvement, the tin- and uranium-containing tantalum-niobium slag in step (1) contains oxides and fluorides.

[0008] As a further improvement, the concentration of the acid leaching agent in step (1) is 0.4-0.8 mol / L.

[0009] As a further improvement, the leaching conditions in step (1) are as follows: liquid-to-solid ratio of 1:0.8-1:1.2, leaching temperature of room temperature, leaching time of 3-5 hours, and stirring during the leaching process at a stirring rate of 300-500 r / min.

[0010] As a further improvement, the amount of flux added in step (2) is 3%-7% of the mass of the tin-enriched slag.

[0011] As a further improvement, the amount of elemental carbon added in step (2) is 15-20% of the mass of the tin-enriched slag.

[0012] As a further improvement, the heat preservation melting temperature in step (2) is 1280-1350℃.

[0013] As a further improvement, the heating rate during step (2) is controlled at 8~10℃ / min, and the holding and melting time is 1.5-2.5h.

[0014] As a further improvement, in step (2), elemental carbon is converted into CO, the system is maintained in a reducing atmosphere, and tin is reduced to elemental form.

[0015] Compared with the prior art, the beneficial effects of the present invention are as follows: This invention employs a two-step combined process of "low-acid wet leaching to enrich tin and high-temperature pyrometallurgical smelting to extract crude tin". Tantalum and niobium slag is leached with low-concentration acid to obtain tin-enriched slag (with Sn content increased by more than 3 times compared to the raw material) and a leachate containing uranium. Crude tin can be obtained by pyrometallurgical smelting of the tin-enriched slag. Uranium in the acid leachate can be recovered by extraction. This invention enables the effective utilization of resources such as tin and uranium in tantalum and niobium slag, with a high recovery rate, a short and simple process, and low cost. At the same time, it reduces the volume and resources of tantalum and niobium tailings, reducing the total amount of waste residue by about 30%. Attached Figure Description

[0016] To more clearly illustrate the technical solutions in the embodiments of the present invention or the prior art, the drawings used in the description of the embodiments or the prior art will be briefly introduced below. Obviously, the drawings described below are some embodiments of the present invention. For those skilled in the art, other drawings can be obtained based on these drawings without creative effort.

[0017] Figure 1 This is a schematic diagram of the process flow of the present invention. Detailed Implementation

[0018] To facilitate understanding of the present invention, the present invention will be described more fully and in detail below with reference to the accompanying drawings and preferred embodiments, but the scope of protection of the present invention is not limited to the following specific embodiments.

[0019] Unless otherwise defined, all technical terms used herein have the same meaning as commonly understood by those skilled in the art. The technical terms used herein are for the purpose of describing particular embodiments only and are not intended to limit the scope of the invention.

[0020] Unless otherwise specified, all raw materials, reagents, instruments and equipment used in this invention can be purchased from the market or prepared by existing methods.

[0021] This invention utilizes a synergistic process of "wet leaching enrichment followed by pyrometallurgical smelting purification" to recover tin and uranium from tantalum-niobium slag. It is particularly suitable for the efficient treatment of tantalum-niobium smelting waste containing tin and uranium, achieving stepwise recovery of tin and uranium and reduction of radioactive waste.

[0022] In some specific embodiments, the wet-pyrometallurgical combined process for recovering tin and uranium from radioactive tantalum-niobium slag of the present invention includes the following steps: (1) Wet leaching enrichment: using tantalum-niobium slag containing tin and uranium as raw material, adding acid leaching agent to the raw material for leaching treatment, controlling the leaching conditions to enrich tin in the solid phase to form tin-enriched slag, and dissolving uranium in the liquid phase to form uranium-containing leachate, and then separating the solid and liquid to obtain tin-enriched slag and uranium-containing leachate.

[0023] In some specific embodiments, the tin- and uranium-containing tantalum-niobium slag is tailings obtained by smelting tantalum-niobium ore using the hydrofluoric acid-sulfuric acid method. Specifically, it is smelting tailings formed by decomposing tantalum-niobium ore using hydrofluoric acid and sulfuric acid and then extracting the slurry.

[0024] In some specific embodiments, the acid leaching agent is at least one acid such as sulfuric acid, hydrochloric acid, or organic acid.

[0025] In some specific embodiments, the leaching conditions include: acid concentration of 0.3-2 mol / L, preferably 0.4-0.8 mol / L, liquid-to-solid ratio of 1:0.8-1:1.2 (mass ratio), leaching temperature of room temperature (20-35℃), leaching time of 3-5 h, and stirring during the leaching process at a stirring rate of 300-500 r / min.

[0026] When the acid concentration in this invention is 0.4-0.8 mol / L, it is considered low-acid leaching, which results in lower costs and less impurity leaching, thus facilitating subsequent uranium extraction. If the acid concentration is too high, it will reduce the tin recovery rate.

[0027] In some specific embodiments, the solid-liquid separation is carried out by suction filtration or pressure filtration, with the filter material having a pore size of 0.22-0.5μm and a separation pressure of 0.1-0.5MPa, to ensure that the tin-enriched slag and the uranium-containing leaching solution are not mixed together.

[0028] In this step, dilute sulfuric acid provides H₂. + As a reaction medium, the soluble components in the raw materials (including some oxides, fluorides, sulfates, of which the fluorides are mainly FeF2, MnF2, etc., and the target element uranium in the form of UF4) undergo dissociation and coordination reactions, and finally achieve dissolution. Typical reactions are shown in formulas (1)-(4). The oxides of elements such as tantalum, niobium, zirconium, and tin do not dissolve in this dilute acid system due to their chemical inertness and remain in the solid phase.

[0029] Fe2O3+6H + =2Fe 3+ +3H2O (1) MnF2+2H + =Mn 2+ +HF (2) FeF2+2H + =Fe 2+ +2HF (3) UF4+4H + +2SO4 2- =U(SO4)2+4HF (4) (2) Pyrometallurgical smelting and purification: Add flux and reducing agent to the tin enrichment slag obtained in step (1), and carry out heat preservation smelting at a preset temperature to reduce and separate tin from the slag phase. After cooling, crude tin product and smelting slag are obtained.

[0030] In some specific embodiments, the co-solvent is selected from at least one of CaF2, NaF, and Na2CO3, and the amount of co-solvent added is 3%-7% of the mass of tin-enriched slag, which can reduce the melting point of the material.

[0031] In some specific embodiments, 15-20% by weight of elemental carbon (e.g., coke) is added as a reducing agent for tin.

[0032] In some specific embodiments, the preset temperature is 1200-1500℃, preferably 1280-1350℃, and the holding and melting time is 1.5-2.5h; In some specific embodiments, an electric furnace is used for heating. During the heating process, the heating rate is controlled at 8~10℃ / min, and the temperature is raised to the preset temperature and then kept warm.

[0033] In some specific embodiments, coke is incompletely combusted to convert into CO, maintaining a reducing atmosphere in the system and reducing tin to elemental form. In actual operation, coke reacts with oxygen in the reactor to generate CO, and the pressure inside and outside the furnace is balanced using the furnace gaps to maintain a reducing atmosphere inside the furnace.

[0034] After wet enrichment in step (1), although tin is directionally enriched, high-melting-point oxides such as zirconium and silicon are also retained and enriched in large quantities along with tin, resulting in a high overall melting point of the tin-enriched slag. Therefore, a fluxing agent needs to be added, as shown in formula (5). The fluxing agent converts the high-melting-point oxides into silicates with lower melting points. Crude tin products are obtained, and the reaction formulas are shown in (6)-(7).

[0035] 2CaF2+SiO2=CaSiO3+2HF (5) SnO2 + 2CO = Sn + 2CO2 (6) SnO + CO = Sn + CO2 (7) Compared with traditional processes for treating tantalum and niobium slag, the method of this invention significantly improves resource utilization while effectively reducing slag volume. The overall process is short, simple to operate, and has good application value.

[0036] Example 1 500g of tantalum-niobium slag (composition shown in Table 1) was leached with a 0.5mol / L sulfuric acid solution with a liquid-to-solid ratio of 1:1 at room temperature for 4 hours with stirring at a stirring rate of 400r / min. After leaching, the residue was separated by filtration to obtain tin-enriched slag and uranium-containing leachate. The Sn grade in the tin-enriched slag was more than 3 times higher than that in the raw material.

[0037] Table 1 Main components of tantalum-niobium slag

[0038] Add 5% CaF2 and 20% coke by mass to tin-enriched slag, control the heating rate at 10℃ / min, and smelt at 1300℃ for 2 hours to obtain crude tin product.

[0039] In this process, the uranium leaching rate reached 91.02%, the total tin recovery rate was 90.48%, and the slag volume was reduced to 70.12% compared to the raw material, achieving efficient synergistic recovery of tin and uranium and reducing slag volume.

[0040] Example 2 Referring to Example 1, 500g of tantalum-niobium slag from the same batch was taken and leached for 4 hours at room temperature using a 0.1mol / L sulfuric acid solution with a liquid-to-solid ratio of 1:1. Other parameters and subsequent pyrometallurgical smelting parameters remained the same as in Example 1.

[0041] The results showed that the uranium leaching rate was only 20.51%; due to the low acid concentration, the tin enrichment was reduced, resulting in a poorer pyrometallurgical effect and a slight decrease in the total tin recovery rate to 80.35%.

[0042] Example 3 Referring to Example 1, 500g of tantalum-niobium slag from the same batch was taken, and wet leaching was completed according to the parameters of Example 1. The pyrometallurgical smelting temperature was then adjusted to 1200℃, while the other conditions remained unchanged.

[0043] The results showed that the uranium leaching rate was 89.76%, while the pyrometallurgical recovery rate of tin plummeted to 60.83%. This was mainly due to the incomplete reaction between the CaF2 additive and the slag phase at low temperatures, and the high melting points of tantalum and niobium oxides, which failed to effectively lower the melting point of tin. 1300℃ is the optimal temperature to ensure complete reduction and separation of tin. These results demonstrate that the tin recovery rate is highly sensitive to smelting temperature.

[0044] Example 4 Referring to Example 1, 500g of tantalum-niobium slag from the same batch was taken. The only difference was the use of a 2mol / L sulfuric acid solution. Other parameters and subsequent pyrometallurgical smelting parameters remained the same as in Example 1.

[0045] The results showed that the uranium leaching rate increased to 93.76%, but the increase was limited. Due to the high acid concentration, the tin slag inflow rate decreased, and the total tin recovery rate dropped slightly to 86.28%.

[0046] Comparative Example 1 Referring to Example 1, 500g of tantalum-niobium slag from the same batch was taken, but only the pyrometallurgical process was used, which was: smelting at 1300℃ for 2 hours.

[0047] The test results showed that the tin recovery rate was 66.84%, far lower than that of Example 1. These results indicate that tantalum-niobium slag without tin enrichment has poor smelting performance and makes subsequent uranium extraction difficult. This comparative example confirms that the combined process of "wet leaching enrichment - pyrometallurgical smelting purification" is key to achieving efficient recovery of tin and uranium from tantalum-niobium slag, and that a single pyrometallurgical process cannot meet the requirements for the synergistic recovery of both elements.

[0048] Comparative Example 2 Referring to Example 1, 500g of tantalum-niobium slag from the same batch was taken and the process parameters were exactly the same as those in Example 1, except that CaF2 flux was not added.

[0049] The results showed a uranium leaching rate of 92.18%, but no tin alloy phase was detected during the pyrometallurgical smelting process. A solidified substance after melting was observed on the surface, while the lower layer consisted of unmelted powder, and no metallic luster was observed. Based on the experimental phenomena and the detection results of the lower unmelted powder, the tin content in the lower unmelted powder was close to that in the raw material, and 9.76% carbon was also present in the lower layer. The reason for this is that the high melting points of oxides of elements such as Ti, Ta, and Nb lead to high material melting temperatures, resulting in poor melting effects at the smelting temperature.

[0050] Comparative Example 3 Referring to Example 1, 500g of tantalum-niobium slag from the same batch was taken and the process parameters were exactly the same as those in Example 1, except that the melting temperature was 1500℃.

[0051] The results showed that the uranium leaching rate was 92.11%, while the tin recovery rate decreased to 78.16%. In production practice, excessively high smelting temperatures have several drawbacks: firstly, they waste energy and shorten the service life of refractory materials; secondly, the excessively vigorous reaction of materials in the furnace leads to intense turbulence of the melt, which is detrimental to the clarification and separation of slag and tin; and thirdly, excessively high tin melt temperatures increase the solubility of impurities in the tin melt, thus affecting product purity. For example, iron can alloy with metallic tin to form an Fe-Sn alloy, which has a high melting point and is easily entrained in the smelting slag during the slag-tin separation process, causing tin loss and ultimately leading to an increase in the tin content in the smelting slag.

[0052] The experimental parameters and results of the above embodiments and comparative examples are shown in the table below.

[0053] Table 2

[0054] The above are merely preferred embodiments of the present invention and are not intended to limit the present invention in any way. Therefore, any simple modifications, equivalent changes, and alterations made to the above embodiments based on the technical essence of the present invention without departing from the scope of the present invention should fall within the protection scope of the present invention.

Claims

1. A process for recovering tin and uranium from radioactive tantalum-niobium slag using a combined wet and pyrometallurgical process, characterized in that, Includes the following steps: (1) Using tin- and uranium-containing tantalum-niobium slag as raw material, acid leaching agent is added to the raw material for leaching treatment, so that tin is enriched in the solid phase to form tin-enriched slag, and uranium is dissolved in the liquid phase to form uranium-containing leachate. Then, solid-liquid separation is performed to obtain tin-enriched slag and uranium-containing leachate. The acid leaching agent is at least one of sulfuric acid, hydrochloric acid, and organic acid, and the concentration of the acid leaching agent is 0.3-2 mol / L; (2) Add flux and reducing agent to the tin enrichment slag obtained in step (1), and heat-hold smelting at a temperature of 1200-1500℃ to reduce and separate tin from the slag phase. After cooling, crude tin product and smelting slag are obtained. The co-solvent is selected from at least one of CaF2, NaF, and Na2CO3; the reducing agent is elemental carbon.

2. The process for recovering tin and uranium from radioactive tantalum-niobium slag using a combined wet and pyrometallurgical process according to claim 1, characterized in that, The tin- and uranium-containing tantalum-niobium slag in step (1) is the tailings obtained by smelting tantalum-niobium ore using the hydrofluoric acid-sulfuric acid method.

3. The process for recovering tin and uranium from radioactive tantalum-niobium slag using a combined hydrometallurgical and pyrometallurgical process according to claim 1, characterized in that, The tin- and uranium-containing tantalum-niobium slag in step (1) contains oxides and fluorides.

4. The process for recovering tin and uranium from radioactive tantalum-niobium slag using a combined hydrometallurgical and pyrometallurgical process according to any one of claims 1 to 3, characterized in that, The concentration of the acid leaching agent in step (1) is 0.4-0.8 mol / L.

5. The process for recovering tin and uranium from radioactive tantalum-niobium slag using a combined hydrometallurgical and pyrometallurgical process according to any one of claims 1 to 3, characterized in that, The leaching conditions in step (1) are as follows: liquid-to-solid ratio of 1:0.8-1:1.2, leaching temperature of room temperature, leaching time of 3-5 hours, and stirring during the leaching process at a stirring rate of 300-500 r / min.

6. The process for recovering tin and uranium from radioactive tantalum-niobium slag using a combined hydrometallurgical and pyrometallurgical process according to any one of claims 1 to 3, characterized in that, The amount of flux added in step (2) is 3%-7% of the mass of the tin-enriched slag.

7. The process for recovering tin and uranium from radioactive tantalum-niobium slag using a combined hydrometallurgical and pyrometallurgical process according to any one of claims 1 to 3, characterized in that, The amount of elemental carbon added in step (2) is 15-20% of the mass of the tin-enriched slag.

8. The process for recovering tin and uranium from radioactive tantalum-niobium slag using a combined hydrometallurgical and pyrometallurgical process according to any one of claims 1 to 3, characterized in that, The heat preservation and melting temperature in step (2) is 1280-1350℃.

9. The process for recovering tin and uranium from radioactive tantalum-niobium slag using a combined hydrometallurgical and pyrometallurgical process according to any one of claims 1 to 3, characterized in that, In step (2), the heating rate is controlled at 8~10℃ / min during the heating process, and the holding and melting time is 1.5-2.5h.

10. The process for recovering tin and uranium from radioactive tantalum-niobium slag using a combined hydrometallurgical and pyrometallurgical process according to any one of claims 1 to 3, characterized in that, In step (2), elemental carbon is converted into CO, the system is maintained in a reducing atmosphere, and tin is reduced to elemental form.

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