WO2022018489A1 - Procedimiento para la lixiviación de elementos de valor a partir de residuos metalúrgicos - Google Patents
Procedimiento para la lixiviación de elementos de valor a partir de residuos metalúrgicos Download PDFInfo
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- WO2022018489A1 WO2022018489A1 PCT/IB2020/056894 IB2020056894W WO2022018489A1 WO 2022018489 A1 WO2022018489 A1 WO 2022018489A1 IB 2020056894 W IB2020056894 W IB 2020056894W WO 2022018489 A1 WO2022018489 A1 WO 2022018489A1
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- leaching
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- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B13/00—Obtaining lead
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- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B13/00—Obtaining lead
- C22B13/06—Refining
- C22B13/08—Separating metals from lead by precipitating, e.g. Parkes process
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- Y—GENERAL TAGGING OF NEW TECHNOLOGICAL DEVELOPMENTS; GENERAL TAGGING OF CROSS-SECTIONAL TECHNOLOGIES SPANNING OVER SEVERAL SECTIONS OF THE IPC; TECHNICAL SUBJECTS COVERED BY FORMER USPC CROSS-REFERENCE ART COLLECTIONS [XRACs] AND DIGESTS
- Y02—TECHNOLOGIES OR APPLICATIONS FOR MITIGATION OR ADAPTATION AGAINST CLIMATE CHANGE
- Y02P—CLIMATE CHANGE MITIGATION TECHNOLOGIES IN THE PRODUCTION OR PROCESSING OF GOODS
- Y02P10/00—Technologies related to metal processing
- Y02P10/20—Recycling
Definitions
- the invention deals with a procedure for the leaching of elements of value from metallurgical waste, in particular, focused on the recovery of elements such as copper and lead, and which may optionally contemplate the leaching of elements such as iron, arsenic, antimony, bismuth, silver and germanium.
- the invention relates to a process for leaching elements of metallurgical waste value to produce a final waste that is a stable waste, in accordance with the TCLP hazard tests (Total characteristic leaching procedure, Leaching procedure for characterization total in Spanish) and SPLP (Synthetic precipitation leaching procedure, Synthetic precipitation leaching procedure in Spanish).
- a process for obtaining a lead concentrate from metallurgical waste is disclosed, which in an even more preferential aspect corresponds to lead carbonate.
- metallurgical waste is dust that comes from a metal smelting process.
- metallurgical residues are powders that come from a copper smelting process.
- metallurgical residues contemplate materials that have already been subjected to leaching processes, such as sulfuric leaching.
- the copper in the sludge is found mainly forming ferrite and/or spinel-type species of the CuFe 2 O 4 form, as well as zinc, ZnFe 2 O 4 and an important part of iron. FeFe 2 O 4.
- the leaching of these species is a function of temperature, acid concentration, and residence time, as described in the study by BS Boyanov, et al. in World Academy of Science, Engineering and Technology, Vol 9, 2015, 1592-1598. who carried out a study on the leaching of synthetic zinc, copper and cadmium ferrites, evaluating the aforementioned variables. The results of this study show that ferrites dissolve better in HCl and H 2 SO 4 , at high temperatures and high acid concentrations.
- copper leaching has an asymptotic behavior with respect to the leaching temperature, which, after a reaction time of 60 min in sulfuric medium, reaches copper leaching yields greater than 90% for the temperature range. temperatures between 85 and 90°C.
- the hydrometallurgical lead recovery route allows working at reduced temperatures, reducing energy consumption, and in turn does not produce sulfur dioxide, which is characterized as a gas that is harmful to the environment.
- the hydrometallurgical route makes use of desulfurization agents such as sodium carbonate, ammonium carbonate, sodium bicarbonate, ammonium bicarbonate, sodium hydroxide, sodium citrate, acetic acid, sodium acetate, among others.
- desulfurization agents such as sodium carbonate, ammonium carbonate, sodium bicarbonate, ammonium bicarbonate, sodium hydroxide, sodium citrate, acetic acid, sodium acetate, among others.
- the goal of these processes is to exchange the sulfate ion for other anions to form insoluble salts.
- lead salts such as lead citrate can be calcined to produce lead oxide (Zárate-Gutiérrez and Lapidus, Hydrometallurgy 144 (2014): 124-128.).
- the mixture of citric acid with sodium citrate is beneficial for the leaching of lead sulfate and subsequent crystallization of lead citrate.
- the solubility product constant of anglesite at 20°C is 6.31 ⁇ 10 -7 , indicating that the solubility of PbSO 4 is very low.
- lead forms a series of soluble complexes.
- citrate complex species are present in solution over the pH range 4.6 to 11.5. At pH below 4.6 the presence of lead sulfate is predominant, while at pH above 11.5 the presence of lead hydroxide is dominant.
- leaching is carried out with a molar ratio of citric acid to lead (II) and (IV) oxide of 1:1 and 4:1 to 20 °C between 15 and 60 min of reaction, reaching leaching efficiencies greater than 99% by weight, obtaining Pb(C 6 H 6 O 7 ) H 2 O as the main species (Sonmez and Kumar, Hydrometallurgy 95, no. 1 -2 (2009), 82-86.).
- Pulp density is another important parameter for lead leaching with citrate solutions.
- citrate solutions Within the range of 10 to 50 g/L of anglesite pulp, leached with a 1 M sodium citrate solution, pH 7 at 600 rpm and 25°C, the highest levels of Lead extraction of 90 to 94% was achieved with a pulp concentration of 10 g/L. The higher the concentration of the pulp, the lower the amount of lead extracted.
- This technology basically comprises treating lead residues that comprise lead (II) oxide, lead (IV) oxide and lead sulfate with a citric acid solution, and which can alternatively be treated in combination with sodium citrate at pHs that vary within the range of 1.4 to 6. It is eventually possible to add hydrogen peroxide in a basic environment as a reducing agent to accelerate the leaching reaction of lead (IV) oxide to produce lead citrate (Sonmez and Kumar, Hydrometallurgy 95, no 1-2 (2009), 82-86.).
- the present invention differs from application W02008056125 A1 in that the pH required for leaching varies between 5.33 and 8.8, where a pH equal to 7 is preferably used. Additionally, the present invention proposes recirculating the citrate solution obtained after a precipitation stage with sodium carbonate, in order to be able to leach metallurgical residue from the sulfuric leaching stage again.
- Another strategy to improve lead citrate crystallization is to increase the molar ratio of citric acid to sodium citrate. For example, at an initial crystallization temperature of 35°C and a molar ratio of citric acid to sodium citrate of 0.92, only 0.42% of the lead was present in the filtrate, while the remaining fraction greater than 99% was crystallized.
- the use of acetic acid to dissolve lead paste has been proposed. Consequently, it is observed that the cooling of the solution accompanied by an increase in the ratio of citric acid and sodium citrate (acidification) is effective in increasing the efficiency of crystallization and obtaining recoveries of lead citrate greater than 99% (US 8323373).
- Patent application AU2009350377A1 discloses a method for recovering lead contained in electrolytic paste from lead batteries, by dissolving lead oxide in H 2 SO 4 in the presence of acetate salts, to obtain lead sulfate soluble in said salts, so as to subsequently add carbonate or hydroxide of the same cation to precipitate lead carbonate/oxycarbonate or lead oxide or hydroxide.
- the process claims the recirculation of the solution containing the acetate salts obtained from the lead precipitation stage in order to dissolve new fresh electrolyte paste.
- Sodium hydroxide precipitation was carried out at 83°C, however, there are no specific details of how to perform sodium carbonate precipitation.
- the application differs from the invention in that citrate salts are used for lead leaching, and in that a sulfate removal stage is not required since the invention contemplates a purge that allows maintaining sulfate concentrations below of the saturation limit.
- Patent US 8568670 discloses a method for producing lead carbonate from slag obtained from a bismuth refining process, where the slag is leached with sodium chloride to obtain lead chloride, which is filtered and neutralized to be added to a solution containing ammonium bicarbonate, under an addition of 2 to 3 times the stoichiometric requirement of the reaction, to precipitate lead carbonate by adjusting the pH between 8 to 11 for a period of between 1 to 2 h.
- the present invention differs from the application in that no chloride salts are used for lead leaching. Additionally, precipitation is performed with sodium carbonate at a pH of between 7 and 8 which is added directly to the lead leaching solution, without requiring previous lead chloride precipitation steps.
- US patent 5,545,805 claims a method of immobilizing lead from materials that contain elements that provide hardness such as calcium and magnesium, where the material that contains lead is contacted with a carbonate associated with alkali metal in a sufficient quantity so that the hardening metals react with the carbonate, and add a polyprotic acid oxyanion.
- the carbonate salt is essentially sodium or potassium carbonate
- the polyprotic acid oxyanion is selected from phosphate, borate, selenate, arsenate, chromate or sulfate, to precipitate lead as an oxosal of said oxyanions.
- the method also discloses that the oxosal may be a carbonate salt, in which case lead carbonate would precipitate.
- the pH of lead precipitation is high, around 12.3.
- the patent application differs from the invention in that oxoanions of polyprotic acids are not required for the precipitation of lead.
- Patent application W02005007904A1 discloses a method for desulfurizing a solid mixture resulting from the rupture of lead batteries containing residues of lead oxides, oxysulfates and sulfates, through contact with ammonium, sodium or potassium carbonate in a molar ratio of between 0.1 and 10% excess over the sulfate concentration, and contact with lanarkite solubilizing substances, among which citric acid and citrates are mentioned, in a molar ratio of carbonate to solvent between 1 and 2, 75.
- the lead residues are desulfurized through the action of the aforementioned solubilizing substances, and later the lead is precipitated through the action of carbonates, where the operating temperature of the process is preferably between 60 and 100 ° C.
- This application differs from the present invention in that the amount of lead carbonate added does not require an addition above the stoichiometric amount necessary to precipitate lead based on the sulfate content of the solution, since the considered purge maintains the levels of sulfate to such a level below saturation that it does not impact the precipitation of lead carbonate.
- the invention object of the present application discloses a method to maximize the leaching of copper and lead that includes sulfuric and citric leaching stages in order to remove the Cu and Pb present in the sludge, to subsequently proceed to alkaline leaching.
- the removal of Cu, Fe and Pb in early stages allows chemically modifying the fluff, leaving the silicon species more labile to leaching as shown in the results obtained in the present application.
- Patent US7329396 describes a process for leaching a metal of value from oxidizing materials, such as a lateritic nickel ore, comprising the step of leaching the ore with a leachant comprising a cationic salt (eg magnesium chloride) and HCI.
- a leachant comprising a cationic salt (eg magnesium chloride) and HCI.
- An additional metal chloride or oxidizer (such as that resulting from the leaching operation) may be added.
- the process comprises recovering a valuable metal from the ore comprising the steps of: leaching the ore with a leachant; separating a metal-rich leachate value from the ore in a first solid-liquid separation; oxidizing and neutralizing the value of the metal-rich leachate thus obtained; and separating a magnesium chloride solution from the leachate thus obtained in a second solid-liquid separation.
- the leaching solution is regenerated from the magnesium chloride solution.
- the regeneration of the leaching solution includes a step of producing magnesium oxide from the magnesium chloride solution.
- a difference of the invention with patent US7329396 is that it is pointed out that a pH above 0.4 is preferred in order to precipitate hematite.
- the precipitation of iron hydroxides is totally unfavorable in the present invention, since the concentration ratio between silver and iron amounts to 0.01 g Ag/g Fe, and, consequently, the precipitation of hydroxides of iron can drag the silver present in solution.
- Patent application CA2820631A1 refers to processes that can be effective for treating various materials that comprise various different metals. These materials are they can be leached with HCl to obtain a leachate and a solid. They can then be separated from each other and a first metal can be isolated from the leachate. Then a second metal can be isolated from the leachate. The first and second metals can each be substantially isolated from the leachate. This can be done by controlling the temperature of the leachate, adjusting the pH, further reacting the leachate with HCl, etc. The metals that can be recovered in the form of metal chlorides can eventually be converted to the corresponding metal oxides, thus allowing HCl recovery.
- the various metals can be chosen from aluminum, iron, zinc, copper, gold, silver, molybdenum, cobalt, magnesium, lithium, manganese, nickel, palladium, platinum, thorium, phosphorus, uranium, titanium, rare earth elements and rare metals. .
- the present invention differs from patent application CA2820631A1 in that the former does not require temperatures above 90°C to efficiently carry out the leaching of metals, unlike the application that requires temperatures above 125°C. Furthermore, the leaching of the aluminum-containing material is carried out with an HCl concentration starting at 18% w/w, while the present invention requires HCl concentrations of less than 140 g/L (or less than 11% w/w). ).
- Figure I shows the process diagram of the method disclosed by the present invention.
- Figure II shows the X-ray diffraction spectrum of the lead concentrate obtained in the application examples.
- Figure III shows the Raman spectrum of the lead concentrate obtained in the application examples.
- the invention describes a process for the leaching of copper and lead, from metallurgical residues of foundry powders that have been subjected to a copper leaching process and that comprise copper, iron, lead, and silicon, and optionally arsenic, antimony and bismuth which maximizes the recovery of copper and lead.
- the invention describes a process for the leaching of copper and lead from metallurgical waste foundry powders that have been subjected to a copper leaching process that includes copper, iron, lead, and silicon, and optionally arsenic, antimony, bismuth, silver, and germanium, and that leaves a final residue that is mainly composed of aluminosilicates and that passes the hazard tests according to with the TCLP assay.
- the invention describes a process for obtaining a lead concentrate from metallurgical waste foundry powders that have been subjected to a copper leaching process and that comprise copper, iron, lead, and silicon, and optionally arsenic, antimony, bismuth, silver and germanium.
- the process of the present invention comprises the following stages: a stage (i) of copper leaching from the metallurgical residue (1), where a first acid leaching solution (2) is used to obtain a first copper-rich leaching solution and iron, and optionally arsenic, antimony and bismuth (3) and a first leached fluff having a reduced content of copper and iron, and optionally reduced in arsenic and enriched in lead and silicon (4), a leaching step (ii) of the first leached fluff (4) where said first leached fluff (4) is processed with a first solution of a salt of a carboxylic acid (5), to obtain a second leached fluff depleted in lead (6) and a second solution of lead-enriched leaching (7), a precipitation stage (iii) where a first base (8) is added to the second lead-enriched leaching solution (7) to obtain a first lead concentrate (9), and a first a lead-depleted precipitation solution (10), a stage (iv
- the metallurgical residue to be processed is powder obtained through a metal smelting process.
- said powder obtained by means of a copper smelting process is foundry powder.
- the metallurgical waste has been subjected to a copper leaching process.
- said metallurgical residue was subjected to leaching with H 2 SO 4 .
- the metallurgical residue to be processed comprises the mineralogical species anglesite, coveline, copper spinels in the form CuOFe 2 O 3 , zinc spinels in the form ZnOFe 2 O 3 , magnetite, iron oxide(lll), pyrite , scorodite, mucovite, kaolinite and lead(ll) sulfate.
- the copper contained in the metallurgical residue is present as copper sulfate, chalcocite, covelin and copper spinels in the CuOFe 2 O 3 form .
- the copper contained in the metallurgical residue is present in at least 50% in the form of copper spinel in the CuOFe 2 O 3 form .
- the silicon contained in the metallurgical residue is present as muscovite and kaolinite.
- the lead contained in the metallurgical residue is present as lead(ll) sulfate, galena or lead(ll) oxide.
- the lead is at least 95% as lead(ll) sulfate.
- the first acid leach solution comprises H 2 SO 4 and/or a refinery effluent.
- step (i) is carried out at a concentration of H 2 SO 4 between 150 and 300 g/L, more preferably at a concentration of H 2 SO 4 of 250 g/L.
- step (i) is carried out at a temperature between 50 and 130°C, more preferably at a temperature of 85°C.
- step (i) is carried out for a time between 3 and 12 hours, more preferably at a residence time of 6 hours.
- step (i) is carried out at a solids concentration of between 5 and 20% w/w, more preferably at a solids concentration of 15% w/w.
- the carboxylic acid salt is sodium citrate.
- the sodium citrate solution has a molar concentration of sodium citrate between 0.5 and 1 M.
- stage (ii) the first leached fluff is fed to the sodium citrate solution in a mass ratio of 1:9.
- step (ii) is carried out at a temperature between 20°C and 60°C, more preferably at 40°C.
- step (ii) is carried out for a residence time of between 1 and 23 h.
- step (ii) is carried out at a pH between 5.3 and 8.8, more preferably at a pH of 7.0.
- step (ii) the corresponding acid of the carboxylic acid salt is added for pH adjustment.
- step (ii) citric acid is added for pH adjustment.
- the first base added to step (iii) is a carbonate salt, selected from one of sodium carbonate, sodium bicarbonate or magnesium carbonate. In an even more preferred option, the first base added to step (iii) is sodium carbonate.
- the sodium carbonate that is added to stage (iii) is carried out in a 1:1 stoichiometric ratio with respect to the lead concentration in the third leaching solution.
- step (iii) the precipitation reaction is carried out at a temperature between 20 and 90°C, more preferably at 70°C.
- step (iii) the precipitation reaction is carried out for a time of 0.5 and 6 h.
- step (iii) the precipitation reaction is carried out at a pH between 6 and 9, more preferably at a pH of 7.5.
- stage (iii) the pH adjustment is carried out with a neutralizer such as sodium hydroxide without considering neutralizers that provide hardness to the solution based on calcium and/or magnesium.
- a neutralizer such as sodium hydroxide
- the first precipitation solution from stage (iii) is recycled to stage (ii) to leach the first leached fluff from stage (i).
- a part of the first lead concentrate is recycled to stage (iii) to act as seed.
- the part of the first lead concentrate that is recirculated corresponds to 30% of the total of said first precipitated solid.
- the recycling ratio of the first lead-depleted precipitation solution to stage (iii) is 90%.
- the first lead-depleted precipitation solution recirculated to step (ii) requires a replacement of sodium citrate.
- the replacement of sodium citrate is a solution that contains sodium citrate in a mass ratio of 0.35:1 with respect to the water used to prepare said solution and is added in such a way as to obtain a pulp that has 10% w/w of solids with respect to the second leached fluff.
- the first lead-depleted precipitation solution recycled to step (ii) requires a pH adjustment to 7.0.
- the pH adjustment in stage (ii) is carried out with a citric acid solution of between 600 and 900 g/L.
- the first lead-depleted precipitation solution obtained from step (iii) does not require a sodium sulfate removal step.
- the first lead concentrate consists of lead carbonate.
- the second base used in the leaching of step (iv) is selected from among Mg(OH) 2 , KOH or NaOH.
- the second base that is added in stage (iv) is added in a ratio of between 5 and 10% w/w with respect to the total mass of the alkaline leaching solution, more preferably in a ratio of 6 .0% w/w.
- the leaching reaction of stage (iv) is carried out at a temperature between 70 and 150°C, more preferably at a temperature of 130°C.
- the leaching reaction of stage (iv) is carried out for a residence time of between 1 to 12 hours, more preferably at a residence time of 3 hours.
- the acid used in the leaching of step (v) is HCl.
- step (v) the HCl is supplied in a concentration that varies between 50 and 140 g/L.
- step (v) a chloride salt is added.
- step (v) the chloride environment is increased by adding magnesium chloride.
- step (v) the chloride salt is supplied in such a way that the chloride concentration is between 140 and 240 g/L.
- step (v) is carried out at a pH between -1.5 and -0.25, preferably within the range -0.73 and -0.65. In a preferred option, step (v) is carried out at a temperature in the range of 40 to 95°C.
- the neutralizing slurry of metal precipitation stage (vi) is selected from those between calcium hydroxide, calcium oxide, calcium carbonate, limestone, dolomitic limestone, magnesium carbonate, magnesium hydroxide or magnesium oxide. magnesium.
- the neutralizer slurry of metal precipitation step (vi) is a magnesium oxide slurry.
- step (vi) is carried out at a temperature between 50 and 95°C.
- the neutralizer slurry added in step (vi) is supplied until a pH of between 3 and 7 is reached.
- step (vi) has a residence time of between 0.5 and 3 hours.
- the fifth solution rich in chloride from step (vi) is sent to a magnesium chloride crystallization process.
- the fifth solution rich in chloride from stage (vi) is recirculated to the fourth silver precipitation stage.
- the sulfuric acid solution of step (vii) has a sulfuric acid concentration of between 60 and 275 g/L.
- the sulfuric acid solution of step (vii) is a sulfuric leaching solution of foundry powders.
- the sulfuric acid solution from stage (vii) is the first leaching solution rich in copper and iron, and optionally arsenic and bismuth from stage (i) whose acidity has been adjusted to between 60 and 275g/L.
- stage (vii) of leaching of the first precipitated solid rich in iron, copper and lead, and optionally arsenic is carried out at a temperature between 50 and 95°C.
- the second lead concentrate is recycled to stage (ii).
- the first leaching solution rich in copper, iron and optionally arsenic is sent to a process for leaching copper from foundry powders.
- the first leaching solution rich in copper, iron and optionally arsenic is sent to an arsenic abatement process.
- the sixth leaching solution rich in copper, iron and optionally arsenic is sent to a process for leaching copper from foundry powders.
- the sixth leaching solution rich in copper, iron and optionally arsenic is sent to an arsenic abatement process.
- the arsenic abatement process is selected from those that contemplate the production of ferric arsenate.
- the arsenic abatement process is a scorodite production process.
- Example 11 A refinery effluent solution was prepared (table 4) to which the concentration of H 2 SO 4 was adjusted to 250 g/L, which was placed in a 5 L glass reactor, where 450 g of fluff that was previously subjected to a copper leaching process to form a pulp with 15% p/p solids. The reactor was stirred at 300 rpm for 6 hours at 85°C. Once the reaction time was over, the pulp was filtered in a kitasate system. The results showed Cu leaching yield of 72.0%, Fe leaching yield of 62.0%, As leaching yield of 71.5%, Zn leaching yield of 57.0%. and a mass loss of 38.5%.
- a solution was prepared with 40 L of water to which 14 kg of sodium citrate were added and the pH was adjusted to 7.0 with a 800 g/L citric acid solution. Once the reagents had dissolved, 6 kg of leached fluff were added according to example 3. The top fluff had a Pb content of 15.4%.
- the leaching was carried out at 20°C and stirred at 1,000 rpm for a period of 9 h. A Pb leaching efficiency of 94% was obtained, obtaining a leached sludge that reduced its mass by 24% with a Pb content of 1.19%.
- a solution was prepared with 2 L of water with a concentration between 323 and 368 g/L of sodium citrate at a pH between 5.3 and 8.8. The pH was adjusted with a 800 g/L citric acid solution. Once the reagents had dissolved, the fluff processed according to example 3 was added in a ratio of between 1.2 and 2.3 g of sodium citrate/g of fluff. The top lint had a Pb content between 15.0 and 15.1%. The leaching was carried out at between 30 and 60°C and agitated between 500 and 700 rpm for a period of between 2 and 4 h. The results are shown in Table 5.
- a solution was prepared with 4.7 L of water to which 1,633 g of sodium citrate were added, adjusting the pH to 7.0 with a 800 g/L citric acid solution. Once the reagents had dissolved, 700 g of leached fluff were added according to example 3. The top fluff had a Pb content of 17.4%. The leaching was carried out at 40°C and stirred at 700 rpm for a period of 3 h. Subsequently, the pulp was filtered, and the filtered solution used to carry out a lead carbonate precipitation. The pH of the solution was adjusted to 7.5 with a 400 g/L NaOH solution, and then an amount equimolar sodium carbonate.
- Precipitation was carried out for 1 h, and subsequently the pulp was filtered, while 90% of the filtered solution from the precipitation stage was used to leach fresh sludge.
- a sodium citrate solution was added in a ratio of sodium citrate: water equal to 0.35:1 fresh to adjust the solids content of the pulp to 10% p/p with respect to the content of erases
- the pH of the solution was adjusted to 7.0 with a citric acid solution of 800 g/L.
- the recirculation of the filtration solution from the precipitation stage was repeated until completing a total of 15 cycles.
- a precipitated solid of lead carbonate was used as seed in the precipitation stage.
- the average leaching efficiency was 94%, with a leached solid with a Pb content of 1.33% on average.
- a Pb precipitate with a Pb content of 76% was obtained.
- the sulfate content reached a maximum value of 96 g/L, which remained constant during the last five cycles of the test.
- Other analytes that were concentrated in the execution of the different leaching cycles were Fe, reaching 4 g/L, B ⁇ reaching 1.5 g/L and K reaching 1.5 g/L.
- a pulp was prepared with a sodium hydroxide solution with a concentration between 5.4 and 8.7% p/p and leached lees subjected to consecutive processes of sulfuric and citric leaching with a solids content between 5.0 and 7.0 %p/p.
- the pulp was placed in a 4 L autoclave and heated at a temperature between 100 and 140°C for a time between 1 and 6 hours at 600 rpm. Once the leaching time was over, the pulp was cooled and filtered in a kitasato system. The results are shown in Table 7.
- a pulp was prepared with 6,230 mL of water to which 420 g of sodium hydroxide and 350 g of leached fluff were added, subjected to consecutive copper and lead leaching processes, to obtain a concentration of 6.0% p/p. NaOH and 5.0% p/p solids.
- the pulp was placed in a 10 L glass reactor and heated at 90°C between 1 and 6 hours and stirred at 900 rpm. Once the leaching time was over, the pulp was cooled and filtered in a kitasato system.
- a solution with an HCl concentration between 54 and 160 g/L and a chloride concentration between 140 and 237 g/L was prepared.
- the chloride concentration was increased by addition of magnesium chloride hexahydrate.
- 180 g of pulp subjected to sulfuric and citric leaching processes were added, and on the other hand, pulp subjected to sulfuric, citric and alkaline leaching processes, such as those described in experiments 1 to 37.
- the pulp was fed to a 5 L glass reactor, heated to 90°C and kept under constant stirring for 6 hours. Once the test was finished, the pulp was filtered in a kitasato system. The results of these tests are presented in Table 9.
- Examples 46 to 52 450 g of PLS obtained from hydrochloric leaching tests were placed in a 600 mL beaker and heated to 25 to 80°C. Neutralization of the hydrochloric leaching solution was carried out using magnesium oxide slurry at 15% by volume until reaching a pH within the range between 3 and 6. Subsequently, the pulp was filtered with 45 ⁇ m filter paper. Table 10. Results examples 46 to 52
- the lead concentrate was recirculated to the citrus leach stage.
- 120 g of lead concentrate obtained from the metal precipitate leaching tests and a sodium citrate solution between 0.5 and 1 M at pH 7 adjusted with citric acid were placed in a 5 L glass reactor.
- the pulp was kept between 20 and 70°C and stirred at 700 rpm for 3 h.
- Lead leaching yield varied between 80 and 82% Table 13. Results examples 58 to 61
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PCT/IB2020/056894 WO2022018489A1 (es) | 2020-07-22 | 2020-07-22 | Procedimiento para la lixiviación de elementos de valor a partir de residuos metalúrgicos |
JP2022509100A JP2023542442A (ja) | 2020-07-22 | 2020-07-22 | 冶金残留物から有益な元素をリーチングするための方法 |
CN202080053498.1A CN114269955A (zh) | 2020-07-22 | 2020-07-22 | 从冶金残留物浸出有价元素的方法 |
PE2021002210A PE20220265A1 (es) | 2020-07-22 | 2020-07-22 | Procedimiento para la lixiviacion de elementos de valor a partir de residuos metalurgicos |
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WO2005007904A1 (en) * | 2003-07-18 | 2005-01-27 | Kandy S. A. | Process for the high yield recovery of lead from spent lead-acid batteries with reduced associated production of slag and gaseous emissions |
WO2008056125A1 (en) * | 2006-11-08 | 2008-05-15 | Cambridge Enterprise Limited | Lead recycling |
WO2011013149A1 (en) * | 2009-07-30 | 2011-02-03 | Millbrook Lead Recycling Technologies Limited | Reclaiming of lead in form of high purity lead compound from recovered electrode paste slime of dismissed lead batteries and/or of lead minerals |
US8568670B2 (en) * | 2009-12-08 | 2013-10-29 | Jiangxi Rare Earth and Rare Metals Tungsten Group Holding Co., Ltd. | Process for producing basic lead carbonate |
CA2820631A1 (en) * | 2012-09-26 | 2014-03-26 | Orbite Aluminae Inc. | Processes for treating various materials |
-
2020
- 2020-07-22 CN CN202080053498.1A patent/CN114269955A/zh active Pending
- 2020-07-22 PE PE2021002210A patent/PE20220265A1/es unknown
- 2020-07-22 CA CA3143384A patent/CA3143384A1/en active Pending
- 2020-07-22 WO PCT/IB2020/056894 patent/WO2022018489A1/es active Application Filing
- 2020-07-22 JP JP2022509100A patent/JP2023542442A/ja active Pending
Patent Citations (5)
Publication number | Priority date | Publication date | Assignee | Title |
---|---|---|---|---|
WO2005007904A1 (en) * | 2003-07-18 | 2005-01-27 | Kandy S. A. | Process for the high yield recovery of lead from spent lead-acid batteries with reduced associated production of slag and gaseous emissions |
WO2008056125A1 (en) * | 2006-11-08 | 2008-05-15 | Cambridge Enterprise Limited | Lead recycling |
WO2011013149A1 (en) * | 2009-07-30 | 2011-02-03 | Millbrook Lead Recycling Technologies Limited | Reclaiming of lead in form of high purity lead compound from recovered electrode paste slime of dismissed lead batteries and/or of lead minerals |
US8568670B2 (en) * | 2009-12-08 | 2013-10-29 | Jiangxi Rare Earth and Rare Metals Tungsten Group Holding Co., Ltd. | Process for producing basic lead carbonate |
CA2820631A1 (en) * | 2012-09-26 | 2014-03-26 | Orbite Aluminae Inc. | Processes for treating various materials |
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CA3143384A1 (en) | 2022-01-22 |
JP2023542442A (ja) | 2023-10-10 |
PE20220265A1 (es) | 2022-02-23 |
CN114269955A (zh) | 2022-04-01 |
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