JPS6362453B2 - - Google Patents

Description

[Detailed description of the invention]

The present invention relates to a method for treating a waste metal catalyst, which uses alumina as a main support and separates and recovers valuable metals such as molybdenum, vanadium, cobalt, and nickel from the waste metal catalyst in a high yield. . Alumina is the main carrier, and molybdenum and
Catalysts supported on one type or a combination of cobalt and nickel are widely used in the organic chemical industry. For example, the molybdenum-based catalyst used in the petroleum desulfurization process contains molybdenum, cobalt, nickel, etc. as catalyst components, and the waste catalyst used for desulfurization contains, in addition to these,
Contains metal components such as vanadium and nickel captured from petroleum distillates. In the past, most of these waste catalysts were discarded, but the separation and recovery of these useful metals has become an essential issue not only from the standpoint of pollution prevention but also from the standpoint of resource conservation and resource reuse. Methods for recovering useful metals as described above from waste catalysts include oxidizing and roasting the waste catalyst and then pressurized leaching with ammonia, leaching with an alkaline solution such as caustic soda or soda carbonate, or leaching with an alkaline solution such as caustic soda or soda carbonate. Some methods have been proposed, such as adding such substances, roasting them with soda, and then leaching them with water. However, the method of pressurizing leaching with ammonia after oxidizing and roasting the waste catalyst requires a large amount of leachate because the solubility of vanadium is very low, and therefore a large amount of leachate with a low concentration of useful metals must be treated. No. In addition, in pressure leaching using caustic soda or soda carbonate and ammonia, the leaching rate of cobalt and nickel is very poor, and in order to recover cobalt and nickel, it is necessary to increase the leaching process using acid again. . Therefore, it is economically disadvantageous in terms of equipment costs and running costs. In general, the carrier of spent catalyst is γ-Al ₂ O ₃ , which has strong activity, so if it is leached with acid after oxidation roasting at low temperature, a large amount of alumina will be eluted, making it difficult to perform the separation and recovery process from other metals. . In order to prevent the elution of alumina, oxidation roasting is performed at high temperatures to create a highly active γ-
It is also possible to convert Al ₂ O ₃ into α-Al ₂ O ₃ , which is stable and less likely to react, before leaching, but in this case, the leaching rate of vanadium, cobalt, nickel, etc. is also reduced. The purpose of the present invention is to provide a method for effectively separating and recovering useful metals such as molybdenum, vanadium, cobalt, and nickel contained in the spent catalyst by eliminating the drawbacks of these conventionally proposed methods. This is what I did. The gist of the method of the present invention to achieve this objective is to oxidize a waste metal catalyst containing alumina as a main support and at least molybdenum and vanadium as valuable metals at a temperature of 750°C or higher in excess air. In the first step of roasting, the obtained burnt ore is subjected to the first stage of leaching under atmospheric pressure using mineral acid, and then the leaching residue is further roasted in an autoclave at 120℃ or higher using mineral acid. The second step is to collect the leachate in which valuable metals have been dissolved by performing the second leaching at high temperature.The obtained liquid is brought into contact with hydrogen sulfide to precipitate the dissolved molybdenum as molybdenum sulfide. The third step is to collect molybdenum sulfide by separating the liquid from which molybdenum sulfide has been separated in the third step (referred to as the recovered liquid). This is a method for treating a waste metal catalyst, comprising the following steps: a fourth step of producing a precipitate containing vanadium by adjusting the amount of the precipitate containing vanadium, and separating the precipitate containing vanadium from the liquid. At this time, if the target is a waste metal catalyst that also contains cobalt, cobalt can be added to the liquid separated in the fourth step by adding an alkaline agent to adjust the pH to 8.5 to 9.5. A fifth step is added in which a cobalt-containing precipitate is formed and this cobalt-containing precipitate is separated from the liquid. Furthermore, when dealing with waste metal catalysts that also contain nickel, add an alkaline agent to the liquid separated in the fourth step to adjust the pH to 8.5 to 9.5 to remove cobalt and nickel. A fifth step is added in which a cobalt- and nickel-containing precipitate is formed and this cobalt- and nickel-containing precipitate is separated from the liquid. The details of the present invention will be explained below. Regarding the first step. First, the waste catalyst on a carrier mainly made of alumina was heated to 750°C.
Sufficient air is introduced at the above temperature to burn off the organic matter and sulfur content adhering to the waste catalyst, and while utilizing the heat of combustion, the metals contained are oxidized and roasted to form oxides and sulfates. The sulfur combustion gas generated in this process is normally SO ₂ , but due to the catalytic action and oxygen partial pressure, most of this is converted to SO ₃ and comes out of the furnace, so it is converted into sulfuric acid in an appropriate absorption tower. A more economical method would be to recover it and use it as mineral acid for use in the leaching process. When a support mainly made of alumina is oxidized and roasted at a temperature of 750℃ or higher, γ-Al ₂ O ₃ changes to α-Al ₂ O ₃ , so even if it is leached with mineral acid, the elution of aluminum will be extremely small. At the same time, the leaching rate of metals such as cobalt and nickel is generally low, but in the leaching process according to the present invention, these metals such as molybdenum, vanadium, cobalt,
Useful metals such as nickel can be forcibly leached out. Regarding the second process. In the second step, valuable metals are forcibly leached from the burnt ore using mineral acid at the same time. As the mineral acid, sulfuric acid, hydrochloric acid, nitric acid, etc. can be used, but sulfuric acid is advantageous for the above-mentioned reasons, so the leaching treatment using sulfuric acid will be described below. Grinding the waste catalyst into suitable particles after oxidizing and roasting improves the leaching rate in the leaching process and also has great benefits in the flow process. In this case, if wet pulverization is performed using process water or the like, dust generation can be prevented. For leaching treatment, use a sulfuric acid solution such that the amount of free sulfuric acid in the liquid after leaching the spent catalyst burnt ore is 50 g/or more, and heat it to 50 to 100°C, preferably 70°C or higher,
Most of the easily soluble molybdenum, vanadium,
Part of the cobalt and nickel is leached (first stage leaching), concentrated and separated in a cone tank or thickener, and the leached liquid is sent to the next useful metal recovery process. Adding sulfuric acid to the Spigot and performing pressure leaching, vanadium and cobalt, which are difficult to leach, are extracted.
Dissolve the nickel (second stage leaching). At this time, the concentration of sulfuric acid as the leaching solution should be 300 to 400 g/L, and the leaching temperature should be 120℃ or higher, preferably 130 to 150℃.
It is preferable that the pressurization conditions be within the range of 1 and the pressure conditions are higher than the water vapor pressure at the leaching temperature. Good results can be obtained if the leaching temperature and pressure are high, but if economic efficiency such as equipment, running costs, etc. and leaching rate are taken into consideration, sufficient results can be obtained under the conditions within the above-mentioned range. Final (second
The reason why the sulfuric acid concentration in the leaching conditions in stage 1) is set to 300 to 400 g/ is that if it exceeds 400 g/, the leaching rate will not improve much, and if it is less than 300 g/, leaching will be insufficient. Also, the temperature should be 120℃ or higher, preferably 130℃.
The reason why the temperature is set at ~150°C is that there is almost no difference from ordinary pressure high temperature leaching below 120°C, and there is no commensurate effect even if the temperature exceeds 150°C. This second-stage leachate is returned to the first-stage leaching process after separating the solids, and is used as the first-stage leachate by adding process water etc. to adjust the acid concentration to that of the first-stage leachate. Can be used repeatedly. Regarding the third step. The obtained leachate is heated to 40℃ or higher, preferably 50 to 80℃.
After heating to ℃, molybdenum in the liquid is separated into molybdenum sulfide by feeding hydrogen sulfide gas. The relationship between reaction temperature and molybdenum recovery rate is shown in Figure 1. It can be seen that molybdenum can be almost completely separated from the liquid at a reaction temperature of 40°C or higher. The reaction between hydrogen sulfide and molybdenum can be thought of as follows. Molybdenum dissolved in sulfuric acid is Mo ^(VI) (SO ₄ ) ₃ , and when it reacts with hydrogen sulfide, it becomes Mo ^(VI) (SO ₄ ) ₃ +3H ₂ S→MoS ₂ +3H ₂ SO ₄ +S …(1) The reaction should generally be completed when the amount of hydrogen sulfide is 3 or more equivalents of molybdenum, but the reaction is likely to be completed due to the VO ₂ ⁺ ions of vanadium dissolved in the liquid (there are various possible ion forms of vanadium, but we will explain with typical examples). ) is thought to react with hydrogen sulfide as shown in the following equation. (V ^(V) O ₂ ) ₂ SO ₄ +2H ₂ S → (V ^(III) O) ₂ S
O ₄ ＋2H ₂ O＋2S...(2) As shown in equation (2), vanadium ions are reduced and hydrogen sulfide is consumed, so in order to recover molybdenum as molybdenum sulfide, the amount of dissolved molybdenum and vanadium and the reaction temperature must be determined. The amount of hydrogen sulfide gas that takes into account the amount of hydrogen sulfide gas dissolved in the liquid is required. Naturally, it is better for the reactor in this case to be airtight, and the recovery of molybdenum sulfide will be further improved if the reaction is carried out under pressure. Note that free sulfur generated when vanadium is reduced increases the S grade of molybdenum sulfide. If it is difficult to obtain hydrogen sulfide gas, a device that emits sodium hydrogen sulfide H ₂ S gas may be used. Regarding the fourth step. In the fourth step, an alkaline agent is added to the liquid obtained in the third step to adjust the pH to a range of 4.5 to 5.5, thereby producing and separating a precipitate containing vanadium (aluminum is also present). At that time,
When sulfuric acid is used as the mineral acid for leaching in the second step, a calcium compound such as calcium carbonate, slaked lime, or quicklime is added in advance to the liquid obtained in the third step and reacted (approximately pH 2 or less). It is advantageous to first separate the free sulfuric acid content as gypsum. If this is omitted, when a precipitate containing vanadium and aluminum is generated, free sulfuric acid will also be neutralized and precipitated, and the vanadium content in the precipitate generated at this time will be low, and the desired vanadium will be removed from this precipitate. Since the extraction process becomes complicated, it is better to separate and remove free sulfuric acid as gypsum in advance. After separating free sulfuric acid as gypsum, one or more of calcium carbonate, slaked lime, and quicklime is further added to the solution to achieve a pH of 4.5 to 5.5, preferably 4.7.
When adjusted to ~5.0 and reacted, a gypsum-like precipitate containing vanadium and aluminum is produced. Figure 2 shows the relationship between the recovery rates of vanadium and aluminum depending on the pH. At this time, neutralization with calcium salts is advantageous because sulfate radicals bonded with vanadium and aluminum form as gypsum, so even if aluminum hydroxide is contained, the permeability of this precipitate is very good, and the process Easy to operate. This point is also one of the characteristics of the method of the present invention. Regarding the fifth step. If the spent catalyst contains cobalt or nickel, the liquid separated in the fourth step is then sent to the fifth step. In the fifth step of producing a precipitate containing cobalt, nickel, etc., one or both of slaked lime and quicklime is added to the separated liquid, the pH is adjusted to 8.5 to 9.5, and the mixture is reacted to contain cobalt and nickel. A gypsum-based precipitate is formed. Figure 2 also shows the relationship between the recovery rates of cobalt and nickel depending on the pH. As in the previous step, by neutralizing with calcium salt, sulfate radicals bonded to cobalt and nickel are formed as gypsum, so even if it contains hydroxides of cobalt and nickel, this precipitate It has very good susceptibility and can be easily operated in the process.
This point is also one of the characteristics of the method of the present invention. As shown in the examples below, even after the hydroxides of nickel and cobalt are separated and recovered in the fifth step, there are still small amounts of each useful metal remaining in the liquid, so the pH was raised to 10 or higher. These useful metals are precipitated all at once and separated, and these precipitates are
It can be completely recovered by returning it to the process and releaching. As described above, according to the present invention, it is possible to recover useful metals at extremely high yields from metal waste catalysts with alumina as the main support, which allows for effective use of resources and pollution caused by the disposal of the waste catalysts. It is also extremely useful for preventing. The present invention will be explained in more detail with reference to Examples below. Example 1 A waste catalyst having the composition shown in Table 1 is oxidized and roasted in a fluidized roasting furnace at an internal temperature of 850 to 900° C. while supplying sufficient air. Table 2 shows the roasted products obtained by oxidative roasting.

【table】

[Table] Next, 650 g of the roasted product was added to a solution 3.3 with a sulfuric acid concentration of 150 g/ml, heated to 90° C., and leached for 1 hour. After the leaching treatment was carried out in this manner, it was allowed to stand still to obtain a supernatant liquid 2. The remaining slurry was transferred to an autoclave, 400 g of sulfuric acid was added, the free sulfuric acid concentration was adjusted to 330 g/L, and the temperature was 150°C.
The leaching residue was washed and dried to weigh 480 g. The composition of this leaching residue is shown in Table 3, and the total leaching rate of the first and second stages is shown in Table 4.

【table】

[Table] The liquid volume of the liquid after the second stage leaching was 1.3, so the above-mentioned residue washing liquid and sulfuric acid were added to this to adjust the liquid volume to 3.3 with a free sulfuric acid concentration of 150g/, and 650g of the same roasted product was obtained. was added, leaching was carried out under the same conditions as in the first stage, and the supernatant liquid No. 2 was obtained by standing still. Table 5 shows the composition of the leachate at this time.

[Table] After heating the leachate 2 shown in Table 5 to 75°C and flowing hydrogen sulfide gas at a flow rate of 1.2/min for 60 minutes,
The hydrogen sulfide gas was stopped, and the reaction was continued for another 30 minutes. The resulting molybdenum sulfide was separated, washed and air-dried, and the composition is shown in Table 6. The weight of the molybdenum sulfide obtained was 155 g.

[Table] The recovery rate of molybdenum as molybdenum sulfide from the leachate was 98.4%. Next, the liquid from which the molybdenum sulfide was separated and the liquid from which the molybdenum sulfide was washed are combined to make 2.9, and the liquid is used to recover vanadium, cobalt, nickel, and aluminum (this liquid is called the recovery liquid). The liquid composition of this recovered liquid is shown in Table 7.

[Table] 130 g of calcium carbonate was added to the above recovered solution 2.9, and after reacting for about 30 minutes, the gypsum was separated. Combine that solution with the cleaning solution used to wash the plaster.
Got 2.9. The combined liquid composition is shown in Table 8.

[Table] Add 330g of calcium carbonate and water to the above combined liquid.
370 ml was added, the pH was adjusted to a range of 4.7 to 5.0, and the reaction was carried out for about 1 hour. Table 9 shows the composition of the precipitate which was separated after the reaction, washed and dried, and the recovery rate of each metal from the recovery solution. The weight of the precipitate obtained at this time was 885 g.

[Table] As shown in Table 9, vanadium and aluminum were almost completely separated by precipitation. Milk of slaked lime was added to solution 6.8, which was a combination of the solution from which vanadium and aluminum were separated as precipitates in the previous operation and the washing solution used to wash the precipitates, the pH was adjusted to 9.0, the reaction was carried out for 1 hour, and the precipitates were removed. generate,
Table 10 shows the composition of the precipitate when it is dried after separation and the recovery rate of each metal from the recovery solution, and Table 11 shows the composition of the combined solution of the solution at that time and the solution used to wash the precipitate.

【table】

[Table] Note that 106 g of the above precipitate was obtained. Also the first
Seven combined liquids having the composition shown in Table 11 were obtained. Next, slaked lime milk was added to composition liquid 7 in Table 11 to adjust the pH to 10.5, and the reaction was carried out for 1 hour to form a precipitate. As a result of analyzing the separated liquid, all of the metal components were detected. Nakatsuta. As described above, vanadium, aluminum, cobalt, and nickel can be recovered and concentrated as precipitates, and each metal can be recovered from these precipitates. Test Example 1 650 g of the roasted product having the composition shown in Table 2 was placed in a 150 g/3.3 sulfuric acid solution, heated to 90°C, leached for 1 hour, left to stand after the leaching treatment, and supernatant liquid 2
The remaining slurry 1.3 was placed in an autoclave, 220 g of sulfuric acid was added, the concentration of free sulfuric acid was adjusted to 246 g/L, the temperature was 150℃, and the pressure was leached at the vapor pressure of the leachate at that temperature for 1 hour, and then separated. ,
The leaching residue was washed and dried to weigh 495 g. Table 12 shows the composition of this leaching residue and the leaching rate under these conditions.

[Table] Test Example 2 650g of the roasted product having the composition shown in Table 2 was placed in a 150g/3.3 sulfuric acid concentration solution, heated to 90°C, leached for 1 hour, left to stand after leaching treatment, and supernatant. liquid 2
The remaining slurry 1.3 was placed in an autoclave, 550 g of sulfuric acid was added, the concentration of free sulfuric acid was adjusted to 425 g/L, the temperature was 150℃, the pressure was the vapor pressure of the leachate at that temperature, and after leaching for 1 hour, it was separated. ,
The leaching residue was washed and dried to weigh 475 g. Table 13 shows the composition of this leaching residue and the leaching rate under these conditions.

[Table] From the above Test Examples 1 and 2, when the concentration of free sulfuric acid in the second stage leachate is 300g/or less, the leaching rate is slightly poor for each metal, and similarly
It was found that even at a free sulfuric acid concentration of more than g/g/g/g/g/g or more, no significant improvement in the leaching rate could be expected. Test Example 3 650 g of the roasted product having the composition shown in Table 2 was placed in a 150 g/3.3 sulfuric acid solution, heated to 90°C, leached for 1 hour, left to stand after the leaching treatment, and supernatant liquid 2
1.3 of the remaining slurry was transferred to an autoclave, 400 g of sulfuric acid was added, and the free sulfuric acid concentration was determined.
After leaching for 1 hour at a temperature of 115° C. and the vapor pressure of the leachate at that temperature, the leaching residue was washed and dried to weigh 490 g. Table 14 shows the composition of this leaching residue and the leaching rate under these conditions.

[Table] Under these conditions, it can be seen that lowering the second stage leaching temperature does not significantly change the leaching rate of each metal component of molybdenum, vanadium, and aluminum, but the leaching rate of cobalt and nickel becomes poor. Test Example 4 After heating 100 ml of the composition liquid shown in Table 5 to 75°C, 10 g of sodium hydrogen sulfide was added, and the mixture was allowed to react for 30 minutes. The generated precipitate was separated and the liquid was analyzed for each metal component, and the recovery rate of each metal component transferred to the liquid is shown in Table 15.

[Table] From the results in Table 15, it can be seen that even without using hydrogen sulfide gas, it is possible to separate molybdenum from other metals using hydrogen sulfide gas, such as sodium hydrogen sulfide.

[Brief explanation of drawings]

Figure 1 shows the relationship between the reaction temperature and the precipitation rate of molybdenum when hydrogen sulfide gas is injected into the leachate to generate and precipitate molybdenum sulfide. FIG. 3 is a diagram showing the relationship between the PH and the residual rate in the liquid when hydroxide is precipitated, and is a process diagram of the present invention.

Claims (1)

[Claims] 1. A first step of oxidizing and roasting a waste metal catalyst containing alumina as a main support and containing at least molybdenum and vanadium as valuable metals at a temperature of 750°C or higher in excess air. After performing the first stage of leaching of the burnt ore using mineral acid under atmospheric pressure, the leaching residue is further subjected to the second stage of leaching using mineral acid in an autoclave at a temperature of 120°C or higher. The second step is to collect the leachate in which valuable metals have been dissolved, and the obtained leachate is brought into contact with hydrogen sulfide to precipitate the dissolved molybdenum as molybdenum sulfide.
The third step is to collect molybdenum sulfide by separating the precipitate.In the third step, an alkali agent is added to the liquid from which molybdenum sulfide has been separated (referred to as the recovered liquid), and the PH4 5 to 5.5 to produce a vanadium-containing precipitate, and a fourth step of separating the vanadium-containing precipitate from the liquid. 2. The first step is to oxidize and roast the waste metal catalyst, which uses alumina as the main support and contains at least molybdenum, vanadium, and cobalt as valuable metals, in excess air at a temperature of 750°C or higher; After performing the first stage of leaching using acid under atmospheric pressure, the leaching residue is further subjected to a second stage of leaching using mineral acid in an autoclave at a temperature of 120°C or higher to extract valuable metals. The second step is to collect the leachate in which the leachate is dissolved, contact the obtained leachate with hydrogen sulfide to precipitate the dissolved molybdenum as molybdenum sulfide,
The third step is to collect molybdenum sulfide by separating the precipitate.In the third step, an alkali agent is added to the liquid from which molybdenum sulfide has been separated (referred to as the recovered liquid), and the PH4 .5 to 5.5 to produce a vanadium-containing precipitate, and this vanadium-containing precipitate is separated from the liquid. A fourth step is to add an alkaline agent to the liquid separated in the fourth step. A method for treating a waste metal catalyst, comprising the steps of: producing a cobalt-containing precipitate by adjusting the pH to 8.5 to 9.5, and separating the cobalt-containing precipitate from the liquid. 3. A first step of oxidizing and roasting a waste metal catalyst containing alumina as a main support and containing at least molybdenum, vanadium, cobalt, and nickel as valuable metals in excess air at a temperature of 750°C or higher, and the resulting burnt ore. After performing the first stage of leaching using mineral acid under atmospheric pressure, the leaching residue is further leached in the second stage using mineral acid at a temperature of 120℃ or higher in an autoclave to obtain a valuable product. The second step is to collect the leachate in which metals have been dissolved, and the obtained leachate is brought into contact with hydrogen sulfide to precipitate the dissolved molybdenum as molybdenum sulfide.
The third step is to collect molybdenum sulfide by separating the precipitate.In the third step, an alkali agent is added to the liquid from which molybdenum sulfide has been separated (referred to as the recovered liquid), and the PH4 .5 to 5.5 to produce a vanadium-containing precipitate, and this vanadium-containing precipitate is separated from the liquid. A fourth step is to add an alkaline agent to the liquid separated in the fourth step. A method for treating a waste metal catalyst, comprising the steps of: producing a cobalt- and nickel-containing precipitate by adjusting the pH to 8.5 to 9.5, and separating the cobalt- and nickel-containing precipitate from a liquid.