Method for synergistically recovering valuable metals from red mud and waste three-way catalyst

Abstract

The invention discloses a method for synergistically recovering valuable metals from red mud and a waste three-way catalyst, and belongs to the technical field of waste resource recycling, the method comprises the following steps: S1, mixing the waste three-way catalyst, the red mud and a slag former, and presintering to obtain a roasted material; s2, the roasted material is crushed, ground, screened, subjected to flash reduction, smelting trapping and atomized powder spraying, and iron alloy containing platinum group metal and water-quenched slag are obtained; s3, the iron alloy containing the platinum group metal is subjected to aqua regia dissolution, nitrate removal and liquid making, and a solution containing the platinum group metal and Fe is obtained; s4, the solution obtained in the step S3 is introduced into an ion exchange column for Fe removal, and a solution containing platinum, palladium and rhodium and an iron-containing solution are obtained; s5, separating the platinum-containing solution, the palladium-containing solution and the rhodium-containing solution to obtain a platinum-containing solution, a palladium-containing solution and a rhodium-containing solution respectively; s6, recovering platinum, palladium and rhodium; according to the method, various metals in the red mud and the waste catalyst can be effectively recovered, the recovery rate is high, and the method has important significance for treating waste with waste.

Description

Technical Field

[0001] The present invention belongs to the technical field of solid waste resource recycling, and particularly relates to a method for synergistically recovering valuable metals from red mud and waste ternary catalysts. Background Art

[0002] Ternary catalysts are generally used in the process of automobile exhaust treatment and can convert CO, HC, NO in automobile exhaust x etc. into CO2, H2O, and N2, etc. However, with the increase of use time, due to inactivation and other factors, automobile ternary catalysts need to be replaced in a timely manner. Moreover, the total precious metal content in waste ternary catalysts is about 2000 g / t, which is about a thousand times the grade of its original ore (2 - 10 g / t), and has great recycling value. At the same time, waste ternary catalysts mainly use cordierite (2MgO·2Al2O3·5SiO2) as the carrier, and the surface is a γ - Al2O3 coating. During their use, some harmful substances such as heavy metals (such as Pb, Cr) and organic substances will be retained. If disposed of randomly, it will cause serious environmental pollution.

[0003] Currently, fire - enrichment and wet - leaching processes are mainly used to recover precious metals from them. However, for the fire - capture recovery of platinum - group metals, the high - temperature reaction time is long, and additional capture agents need to be added, resulting in high costs; for the direct wet treatment of waste catalysts, the leaching rate of platinum - group metals is not high, and a large amount of oxidants need to be used, generating a large amount of wastewater.

[0004] Red mud is a strongly alkaline solid waste residue generated during the production of alumina. It contains a large amount of iron oxide and appears red, so it is called red mud. Due to the strong alkalinity, fine particle size, and complex composition of red mud, especially the high content of Na2O, it is difficult to be widely used in large - scale building materials fields such as cement and concrete. Currently, the comprehensive utilization rate of red mud is less than 10%, and a large amount of red mud needs to be treated by stacking. If the red mud generated by the alumina production process is not effectively utilized, a large amount of industrial solid waste will be generated, which will have an impact on the environment and no economic benefits can be obtained.

[0005] Currently, the methods for extracting iron from red mud mainly include magnetic separation for iron extraction or the pretreatment method of reduction roasting + magnetic separation for iron extraction. Among them, the magnetic separation for iron extraction process has problems such as low recovery rate, low iron grade of the product, and the tailings still need to be landfilled. The reduction roasting + magnetic separation for iron extraction process has problems such as high cost and difficulty in utilizing the magnetic separation tailings. Summary of the Invention

[0006] To solve the above - mentioned technical problems, the present invention provides a method for synergistically recovering valuable metals from red mud and waste ternary catalysts. This method can effectively recover Fe, Pt, Pd, and Rh in red mud and waste catalysts and maintain a high recovery rate.

[0007] The technical solution adopted by the present invention is as follows: A method for synergistically recovering valuable metals from red mud and waste ternary catalysts, the method comprising the following steps: S1: Mix the waste ternary catalyst, red mud, and slag-forming agent, and perform pre-sintering to obtain a calcined material; S2: Crush, grind, and screen the calcined material, and then perform flash reduction, smelting capture, and atomized powder spraying to obtain a ferroalloy containing platinum group metals and water-quenched slag; S3: Dissolve the ferroalloy containing platinum group metals in aqua regia, remove nitrates, and prepare a solution to obtain a solution containing platinum group metals and Fe; S4: Pass the solution obtained in step S3 through an ion exchange column containing 001*7 (732) resin to remove Fe, and obtain a solution containing platinum, palladium, and rhodium and an iron-containing solution; S5: Separate the solution containing platinum, palladium, and rhodium by extraction to obtain a platinum-containing solution, a palladium-containing solution, and a rhodium-containing solution respectively; S6: Recover platinum, palladium, and rhodium from the platinum-containing solution, palladium-containing solution, and rhodium-containing solution respectively.

[0008] In step S1, the mass ratio of the waste ternary catalyst, red mud, and slag-forming agent is (10-40): (45~85): (5~10).

[0009] In step S1, the slag-forming agent is one or more of Cao, CaF2, and CaCO3.

[0010] In step S1, the conditions for pre-sintering are heating and holding at 650~750°C for 0.5-3 h.

[0011] In step S2, the flash reduction temperature is 1073-1473°C, the reduction time is 1-10 s, the flash reduction atmosphere is a mixture of N2 and H2, a mixture of N2 and CO, a mixture of N2 and CH4, or a combination of the three atmospheres, where the volume fractions of H2, CO, and CH4 are 5%-10%; the smelting capture temperature is 1450-1600°C, and the smelting capture time is 0.5-3 h.

[0012] In step S3, the solid-liquid ratio of the ferroalloy containing platinum group metals to aqua regia is 1:1-1:10, and the reaction time is 1-8 h.

[0013] In step S3, removing nitrates means heating and boiling the iron alloy solution containing platinum, palladium, and rhodium, and continuously adding concentrated hydrochloric acid when it is boiled to a small volume to remove nitrates. Concentrated hydrochloric acid is repeatedly added until there is no yellow smoke in the solution, and finally it is evaporated to dryness over a low fire; preparing a solution means dissolving the above product in a hydrochloric acid solution.

[0014] In step S4, after all the solution is added to the ion exchange column, deionized water is added for washing. The washing water is combined with the effluent to form a solution containing platinum, palladium, and rhodium. The resin after washing with deionized water is desorbed with hydrochloric acid 2-3 times to obtain a solution containing FeCl3.

[0015] The solution containing FeCl3 is adjusted to a pH of 4.5, and phosphate is added for precipitation to obtain an iron phosphate product.

[0016] The hydrochloric acid concentration used for the hydrochloric acid desorption is 1-6 mol / L.

[0017] In step S5, the extractants are amine extractants LIX84-I and TOA-TBP, and the extraction sequence is: palladium, platinum, and the remaining solution is a solution containing rhodium.

[0018] In step S6, platinum, palladium, and rhodium can be recovered from the platinum-containing solution, palladium-containing solution, and rhodium-containing solution by any well-known or disclosed method in the prior art. It can also be carried out according to the following method: the platinum-containing solution is obtained by ammonium salt precipitation, calcination, and pickling to obtain sponge platinum; the palladium-containing solution is obtained by ammonia precipitation and hydrazine hydrate reduction to obtain sponge palladium; the rhodium-containing solution is obtained by formic acid reduction and H2 reduction roasting to obtain sponge rhodium powder.

[0019] Compared with the prior art, the present invention has the following beneficial effects: The present invention utilizes the co-disposal of red mud and waste ternary catalyst to co-recover iron and platinum group metals in red mud. In this method, red mud can capture the metals in the waste ternary catalyst, and the Fe2O3 in red mud is pre-reduced synchronously by pre-roasting to reduce the subsequent flash reduction pressure. Then, the roasted material is subjected to flash reduction in a reducing atmosphere and then smelted and captured to obtain an iron alloy containing platinum group metals, and then iron, platinum, palladium, and rhodium are recovered through subsequent refining.

[0020] The method for co-recovering valuable metals from red mud and waste ternary catalyst provided by the present invention can effectively avoid the problem that the high content of Na2O in red mud leads to its inability to be used in the fields of cement and concrete. At the same time, the process route of pyrometallurgical capture + hydrometallurgical enrichment and separation can effectively avoid the problems of additional capture agents added during the recovery of waste catalysts and excessive wastewater and acid gas generated during the direct wet treatment of waste catalysts, which is of great significance for realizing waste treatment with waste. Description of the Drawings

[0021] Figure 1 It is a process flow diagram of the method for co-recovering valuable metals from red mud and waste ternary catalyst in the embodiment of the present invention. Detailed Description of the Invention

[0022] The present invention will be described in detail below with reference to the embodiments.

[0023] Example 1

[0024] Figure 1 It is a schematic diagram of the method for synergistically recovering valuable metals from red mud and waste ternary catalysts in the embodiments of the present invention.

[0025] As Figure 1 shown, the method for synergistically recovering valuable metals from red mud and waste ternary catalysts in this embodiment includes the following steps: S1. Mix the waste ternary catalyst, red mud, and slag-forming agent in a mass ratio of 40:50:10, and pre-roast at 700 °C for 1 h after mixing to obtain a roasted material; S2: After the roasted material is crushed, ground, and screened, it undergoes flash reduction, smelting capture, and atomization to produce platinum group metal-containing ferroalloy powder and water-quenched slag, and the water-quenched slag can be used in the building materials and cement industries; The specific process is as follows: The roasted material is crushed and ground, passed through a 200-mesh sieve to obtain waste roasted material powder with a particle size <74 μm, then flash-reduced at 1373 °C for 9 s. The atmosphere for flash reduction is a mixed gas of N2 and H2, where the volume fraction of H2 is 5%, and then smelted and captured for 180 min at a smelting capture temperature of 1550 °C; S3: Mix the platinum group metal-containing ferroalloy with aqua regia at a solid-liquid ratio of 1:10, place it at 80 °C for oxidative dissolution for 8 h to obtain a solution containing Fe, Pt, Pd, and Rh; then heat and boil until it reaches a small volume, continuously add commercially available concentrated hydrochloric acid to drive off nitrates when the volume is small, repeatedly add concentrated hydrochloric acid until no yellow fumes are generated in the solution, finally evaporate to dryness over a gentle fire, and then dissolve and prepare a solution of the above product with 1 mol / L hydrochloric acid solution; S4: Pass the above solution through an ion exchange column containing 001*7 (732) resin to remove Fe, obtaining a solution containing platinum, palladium, and rhodium and an iron-containing solution. The specific process is as follows: Pass the above solution through an ion exchange column containing 001*7 (732) resin. After all the solution is exchanged, add deionized water for washing, and the washing water and the exchanged solution are combined to form a solution containing Pt, Pd, and Rh; the resin after deionized water washing is eluted 3 times with 6 mol / L hydrochloric acid to obtain an FeCl3 solution, adjust its pH to 4.5, and add phosphate for precipitation to obtain an iron phosphate product.

[0026] S5: Use a LIX84-I type extractant to extract palladium from the obtained solution containing platinum, palladium, and rhodium through single-stage extraction to obtain the organic phase of palladium. After washing and three-stage hydrochloric acid back-extraction, a solution containing palladium is obtained. Subsequently, sponge palladium is obtained through refining in step S6; adjust the acidity of the raffinate after palladium extraction, use TOA-TBP for single-stage extraction of platinum, and obtain a platinum-containing solution after three-stage hydrochloric acid back-extraction. Subsequently, sponge platinum is obtained through refining in step S6; a platinum-containing solution is obtained after three-stage back-extraction of the raffinate; the raffinate after platinum extraction contains rhodium, and it is continuously refined in step S6 subsequently.

[0027] S6: Recover platinum, palladium, and rhodium from the platinum-containing solution, palladium-containing solution, and rhodium-containing solution respectively. The specific process is as follows: ① Heat the platinum-containing solution to above 70 °C, gradually add NH4Cl until no precipitate is formed, and then filter to obtain the filter residue (ammonium chloroplatinate) and the filtrate; add the filter residue to aqua regia for reaction, the reaction temperature is 85 - 95 °C, react for 2 h, filter to obtain the filtrate, drive off the nitric acid in the filtrate 2 - 3 times, then add a certain amount of 1% dilute hydrochloric acid and boil to dissolve, and filter after the solution cools to obtain the filtrate; heat the filtrate to above 70 °C, gradually add NH4Cl until no precipitate is formed, and then filter to obtain the filter residue (ammonium chloroplatinate) and the filtrate. Dry the filter residue and then calcine it in a muffle furnace. Control the initial temperature at 220 - 250 °C, keep it at a constant temperature for 2 h, then raise the temperature to 400 - 450 °C, keep it at a constant temperature for 1 - 2 h, and then raise the temperature to 750 °C, keep it at a constant temperature for 2 - 3 h to obtain sponge platinum; ② Slowly add ammonium chloride to the palladium-containing solution until no red precipitate is formed. After the reaction, filter to obtain the filter residue and the filtrate. Wash the filter residue 3 times with a saturated ammonium chloride solution to obtain pure ammonium chloropalladate; slurry ammonium chloropalladate with 5% water, then add ammonia water (30%) to adjust the pH to 8 - 9, react for 1 - 2 h, then cool and filter. Add hydrochloric acid to the filtrate until the end point pH is 1, filter to obtain the filtrate and the filter residue. Wash the filter residue with 1% dilute hydrochloric acid to obtain pure dichlorodiamminepalladium complex; slurry the pure dichlorodiamminepalladium complex with water, slowly add hydrazine hydrate for reduction to obtain palladium powder (react according to adding 1 L of 50% hydrazine hydrate for every 1 kg of palladium). The palladium powder is washed with water and dried to obtain sponge palladium powder; ③The rhodium-containing solution is heated and concentrated (to an anhydrous state), then dissolved in water, and after filtration (to remove organic impurities), the filtrate is obtained. Hydrochloric acid is added according to the ratio of filtrate:hydrochloric acid = 4:1 and the temperature is raised to 80 °C. 6% by volume of EDTA is added to precipitate rhodium. After reacting for 1 - 2 h, rhodium salts are precipitated to obtain filter residue and filtrate; the filter residue is dissolved in aqua regia for the second time, and after dissolution and nitric acid expulsion operations, filter residue and filtrate are obtained after filtration; a certain amount of sodium formate solution (a neutral solution prepared from formic acid and sodium hydroxide solution) is added to the filtrate in a ratio of 1:10, and then sodium hydroxide solid is added to adjust the pH value of the solution to 13 - 14. After reacting for a period of time, rhodium black and waste liquid are obtained by filtration; the rhodium black is dried and placed in a tubular furnace, and H2 is introduced for reduction. The reduction temperature is 700 °C, and it is heated for 3 h. After cooling, the reduced product, rhodium powder, is obtained.

[0028] Example 2

[0029] In this example, the mass ratio of the waste ternary catalyst, red mud, and slag-forming agent in step S1 of Example 1 is changed from 40:50:10 to 50:40:10, and other conditions are exactly the same, and iron phosphate and sponge platinum, sponge palladium, and sponge rhodium are obtained respectively.

[0030] Example 3

[0031] In this example, the flash reduction time in step S2 of Example 1 is changed to 2 s, and other conditions are exactly the same, and iron phosphate and sponge platinum, sponge palladium, and sponge rhodium are obtained respectively.

[0032] Example 4

[0033] In this example, the flash temperature in step S2 of Example 1 is changed to 1173 °C, and other conditions are exactly the same, and iron phosphate and sponge platinum, sponge palladium, and sponge rhodium are obtained respectively.

[0034] Example 5

[0035] In this example, the smelting and trapping temperature in step S2 of Example 1 is changed to 1500 °C, and other conditions are exactly the same, and iron phosphate and sponge platinum, sponge palladium, and sponge rhodium are obtained respectively.

[0036] Example 6

[0037] In this example, the smelting and trapping time in step S2 of Example 1 is changed to 30 min, and other conditions are exactly the same, and iron phosphate and sponge platinum, sponge palladium, and sponge rhodium are obtained respectively.

[0038] Example 7

[0039] In this example, the solid-liquid ratio of aqua regia to the platinum group metal-containing iron alloy powder in step S3 of Example 1 is changed to 1:3, and other conditions are exactly the same, and iron phosphate and sponge platinum, sponge palladium, and sponge rhodium are obtained respectively.

[0040] Example 8

[0041] In this example, the reaction temperature of aqua regia and the platinum group metal-containing iron alloy powder in step S3 of Example 1 was changed to 50 °C, and other conditions were exactly the same, obtaining iron phosphate, sponge platinum, sponge palladium, and sponge rhodium respectively.

[0042] Example 9

[0043] In this example, the reaction time of aqua regia and the platinum group metal-containing iron alloy powder in step S3 of Example 1 was changed to 2 h, and other conditions were exactly the same, obtaining iron phosphate, sponge platinum, sponge palladium, and sponge rhodium respectively.

[0044] Example 10

[0045] In this example, the hydrochloric acid concentration during the liquid-making process in step S3 of Example 1 was changed to 3 mol / L, and other conditions were exactly the same, obtaining iron phosphate, sponge platinum, sponge palladium, and sponge rhodium respectively.

[0046] Example 11

[0047] In this example, the hydrochloric acid concentration during the salting-out process in step S4 of Example 1 was changed to 1 mol / L hydrochloric acid, and other conditions were exactly the same, obtaining iron phosphate, sponge platinum, sponge palladium, and sponge rhodium respectively.

[0048] Example 12

[0049] In this example, the pickling times during the salting-out process in step S4 of Example 1 were changed to 1 time, and other conditions were exactly the same, obtaining iron phosphate, sponge platinum, sponge palladium, and sponge rhodium respectively.

[0050] The main components of the waste ternary catalyst and red mud used in Examples 1-12 are shown in Tables 1 and 2.

[0051] .

[0052] The recovery rates of Fe, Pt, Pd, and Rh are calculated according to the following formula Recovery rate of platinum group metals = ; Recovery rate of iron =

[0053] The recovery rates of valuable metals in the above Examples 1-12 were calculated as shown in Table 3 .

[0054] According to the above data, it can be seen that the method for synergistically recovering valuable metals from red mud and waste ternary catalyst provided by the present invention can effectively recover Fe, Pt, Pd, and Rh in red mud and waste catalyst, and can maintain a relatively high recovery rate.

[0055] According to the above data, the flash reduction time, smelting and trapping temperature, and smelting and trapping time have a great influence on the recovery rate of platinum group metals. Under relatively economic conditions, the flash reduction time and smelting and trapping time should be extended and the smelting and trapping temperature should be increased. At the same time, the solid-liquid ratio and acid digestion time should be appropriately increased during the acid digestion process; According to the above data, the reduction time and solid-liquid ratio have a certain influence on the iron recovery rate. The reduction time and solid-liquid ratio should be appropriately increased. At the same time, the number of salting-out operations should be increased to ensure the recovery rate of Fe.

[0056] The detailed description of a method for co-recovering valuable metals from red mud and waste ternary catalysts with reference to the embodiments is illustrative rather than restrictive. Several embodiments can be enumerated within the defined scope. Therefore, changes and modifications without departing from the general concept of the present invention shall fall within the protection scope of the present invention.

Claims

1. A method for synergistically recovering valuable metals from red mud and waste ternary catalysts, characterized in that, The method includes the following steps: S1: Mix the waste ternary catalyst, red mud, and slag-forming agent, and perform pre-sintering to obtain a calcined material; S2: Crush, grind, and screen the calcined material, and then perform flash reduction, smelting capture, and atomized powder spraying to obtain a ferroalloy containing platinum group metals and water-quenched slag; S3: Dissolve the ferroalloy containing platinum group metals in aqua regia, remove nitrate, and prepare a solution to obtain a solution containing platinum group metals and Fe; S4: Pass the solution obtained in step S3 through an ion exchange column containing 001*7 (732) resin to remove Fe, and obtain a solution containing platinum, palladium, and rhodium and an iron-containing solution; S5: Separate the solution containing platinum, palladium, and rhodium by extraction to obtain a platinum-containing solution, a palladium-containing solution, and a rhodium-containing solution respectively; S6: Recover platinum, palladium, and rhodium from the platinum-containing solution, palladium-containing solution, and rhodium-containing solution respectively.

2. The method for synergistically recovering valuable metals from red mud and waste ternary catalysts according to claim 1, characterized in that, In step S1, the mass ratio of the waste ternary catalyst, red mud, and slag-forming agent is (10-40):(45~85):(5~10).

3. The method for co-recovering valuable metals from red mud and spent ternary catalysts according to claim 1, wherein In step S1, the slag-forming agent is one or more of Cao, CaF2, and CaCO3.

4. The method for co-recovering valuable metals from red mud and spent ternary catalysts according to claim 1, characterized in that, In step S1, the conditions for pre-sintering are heating and holding at 650~750°C for 0.5-3h.

5. The method for co-recovering valuable metals from red mud and spent ternary catalysts according to claim 1, wherein In step S2, the flash reduction temperature is 1073-1473°C, the reduction time is 1-10s, the flash reduction atmosphere is a mixture of N2 and H2, a mixture of N2 and CO, a mixture of N2 and CH4, or a combination of the three atmospheres, where the volume fractions of H2, CO, and CH4 are 5%-10%; the smelting capture temperature is 1450-1600°C, and the smelting capture time is 0.5-3h.

6. The method for co-recovering valuable metals from red mud and waste ternary catalysts according to claim 1, characterized in that, In step S3, the solid-liquid ratio of the ferroalloy containing platinum group metals to aqua regia is 1:1-1:10, and the reaction time is 1-8h.

7. The method for co-recovering valuable metals from red mud and waste ternary catalysts according to claim 1, wherein In step S3, removing nitrate means heating and boiling the iron alloy solution containing platinum, palladium, and rhodium, adding concentrated hydrochloric acid continuously when it is boiled to a small volume to remove nitrate, adding concentrated hydrochloric acid repeatedly until there is no yellow smoke in the solution, and finally evaporating to dryness over a gentle fire; preparing a solution means dissolving the above product in hydrochloric acid solution.

8. The method for co-recovering valuable metals from red mud and waste ternary catalysts according to claim 1, characterized in that, In step S4, after all the solution is added to the ion exchange column, deionized water is added for washing, and the washing water is combined with the effluent to form a solution containing platinum, palladium, and rhodium; the resin after deionized water washing is analyzed with hydrochloric acid 2-3 times to obtain a solution containing FeCl3.

9. The method for co-recovering valuable metals from red mud and waste ternary catalysts according to claim 8, wherein, The solution containing FeCl3 is adjusted to a solution pH of 4.5, and iron phosphate products are obtained by adding phosphate precipitation.

10. The method for co-recovering valuable metals from red mud and waste ternary catalysts according to claim 8, wherein The hydrochloric acid concentration used for the hydrochloric acid analysis is 1-6mol / L.

Images