Method for treating electronic waste and recovering valuable metals and rare earth elements therefrom

Abstract

A process for treating e-waste, including at least one of used batteries and electronic waste, is provided. The process may include the following steps: fragmenting the e-waste to produce fragmented waste containing metal powder; recovering the powder from the fragmented waste, where the metal powder may include rare earth elements (REEs) and base metals; leaching the REEs and base metals from the metal powder to produce a residual metal-depleted solids, a concentrated leachate, and a secondary leachate; extracting the REEs from the REE-enriched leachate and the secondary leachate; and selectively extracting base metals from the base-metal-enriched leachate to produce a plurality of extracted components, each extracted component including at least one of the base metals. Depending on the composition of the e-waste being treated, precious metals may also be recovered from the e-waste.

Description

[Technical Field]

[0001] The technical field relates generally to the treatment of electronic waste, including at least one of used batteries and electronic waste, and more particularly to process implementations involving selective leaching and extraction of valuable metals and rare earth elements. [Background technology]

[0002] In North America, the management of battery waste and electronic waste is becoming an increasing environmental and political concern due to the increasing volume of waste being managed and the environmental risks associated with the presence of heavy metals (e.g., Cd, Zn, Mn, Ni) in such waste. Battery waste collected within collection centers includes various types of batteries, primarily alkaline batteries (Zn-Mn and Zn-C), nickel-cadmium (Ni-Cd) batteries, nickel-metal hydride (Ni-MH) batteries, lithium (Li) batteries, and sealed lead (Pb) acid batteries.

[0003] Over 26,000 tons of used batteries were collected in North America from 2017 to 2020. However, the majority of battery waste still ends up in landfills or incineration centers, even though this waste contains significant amounts of potentially recyclable metals and rare earth elements (REEs).

[0004] Table 1 shows the content of various metals found in battery waste collected by used battery collection centers according to battery type, as well as the average percentage of different battery types contained in the waste.

[0005] [Table 1]

[0006] In recent years, many technologies have been developed that allow for the treatment of different types of batteries, some of which are now commercialized: Accurec, American Manganese, Batrec, Ecopilas, Inmetco, Recupyl, Retriev Technologies, Revatech, RMC, Sab Nife, Snam, Solvay, and Umicore are examples of companies utilizing treatment processes on an industrial scale.

[0007] Since the mid-1990s, e-waste has been recognized as the fastest-growing component of the solid waste stream. Today, it is experiencing an annual growth rate of 3% to 5%. Thus, e-waste reached over 50 million metric tons (Mt) in 2018. Solving the E-Waste Problem (StEP) estimated that the amount of e-waste generated in the United States in 2018 was nearly 7 Mt, representing an average of 11.6 kg of e-waste per person in the Americas. A 2017 United Nations (UN) Academic Branch Report estimated the value of materials contained in e-waste discarded or collected in 2016 at over US$60 billion. Although e-waste contains certain hazardous components, it also consists of precious metals (Ag, Au, Pd, Pt), strategic metals (Co, In, Li, Mo, Sn), basic metals (Al, Cd, Cr, Cu, Fe, Mn, Ni, Pb, Zn), and rare earth elements (ETR - lanthanides, in addition to Sc and Y), the availability of which is limited on the international market. Despite the potential revenue, only 15% of e-waste is recycled.

[0008] Approximately 20 patents relating to the recycling of waste batteries by hydrometallurgical and pyrometallurgical processes are listed in Table 2. Low recovery rates of Zn and Mn, non-recovery of other metals and rare earth elements, and unrealistic operating conditions are factors that significantly limit the industrial application of some of the processes disclosed in these patents. Indeed, there is a need for better alternatives for the treatment of battery waste that address these issues.

[0009] U.S. Patent No. 8,728,419 (B1) discloses a process for recycling used alkaline batteries. These batteries are primarily made from a steel-cased battery; an alkaline electrolyte; and a mixture of manganese oxide, zinc hydroxide, zinc oxide, and some carbon. In this process, only a small portion of the manganese is soluble in a sulfuric acid solution maintained at a temperature of 60-80°C, while almost all of the zinc is soluble. The resulting slurry is then filtered to obtain a cake containing MnO2 and a leachate containing Mn, Zn, and Fe. Iron is removed from the leachate by heating and air oxidation at pH 4. Soluble MnSO4 found in the leachate is removed as insoluble MnO2 by adding sodium persulfate at pH 4. The high-purity solution of ZnSO4 is then treated by precipitation with Na2CO3 at pH 10-11. ZnCO3 is obtained as the final product. The MnO2 contained in the cake is mixed with H2SO4 and sodium metabisulfite or sulfur dioxide to dissolve the Mn(IV) at 60°C. The pH of this solution is then adjusted to 4, and sodium persulfate is added to form a precipitate of gamma manganese dioxide. While Mn and Zn are effectively recovered, this process is only intended for the treatment of used alkaline batteries.

[0010] A key technology is the Batenus process, disclosed in U.S. Patent No. 5,575,907(A). This patent describes a process used to recycle metals from unsorted used batteries. The primary metals present in the mixture are Mn, Zn, Ni, Cd, Pb, and Hg. Initially, the used batteries are simply processed by mechanical methods to separate the waste into two fractions (coarse and fine). Then, wet chemical processes are used to separately recover each metal. The fine fraction is almost completely leached during two leaching steps, one in the presence of water (first leaching step) and the other in the presence of dilute sulfuric acid and sulfur dioxide (second leaching step). Two cationic exchange resins are then used to remove Hg from the acidic leachate and recover Cu. Zn is then extracted by a liquid-liquid extraction process using an organic extractant. The Cu-, Hg-, and Zn-free solution is then treated by a multi-stage ion exchange process to separate Ni and Cd. Finally, the Hg-, Cu-, Zn-, Cd-, and Ni-free solution is electrolyzed to recover solid MnO2 by pH adjustment. Cu, Cd, Zn, and Ni can also be recovered by electrowinning to obtain the final product in metallic form. However, this process is not designed for the recovery of rare earth elements, cobalt, and lithium.

[0011] European Patent No. 0,620,607 (B1) discloses a process for recovering metals from a mixture of used batteries. The mixture may contain Zn, Mn, Ni, Cu, and Cd in various concentrations. This recycling method focuses on the recovery of these two metals, given the attractive market for these metals. Used batteries are crushed under a stream of cold, dry air, and the ferrous materials are removed from the non-ferrous metals (Hg, Mn, Zn, Cd, and Ni) using a magnetic separation process. Inert materials are then separated from the mineral sludge by flotation. The mineral sludge is then treated by leaching with H2SO4 at temperatures between 40 and 90°C in the presence of a reducing agent. Cu is then recovered from the leachate by cementation. Ni and Cd are selectively electrodeposited at a pH of 4.0 to 5.5 using a potential between 1.5 and 5.0 V. Finally, Zn and Mn are simultaneously recovered using an electrowinning process.

[0012] Two other processes have also been developed for the treatment of alkaline batteries (European Patent No. 1,454,376 B1) and lithium-based batteries (U.S. Patent No. 7,820,317 B2). Alkaline batteries are treated with ultrasonically accelerated sulfuric acid leaching in the presence of a reducing agent. After removing other heavy metal impurities by high-temperature carburization, manganese is recovered as manganese carbonate. Zinc is separated as a soluble zinc-ammonia complex before finally being subjected to heating to obtain the zinc salt. The second patent concerns lithium batteries. The soluble lithium obtained after several processing steps is suspended in water by adding LiOH, thereby increasing the pH of the solution to 12 before hydrolysis. Lithium can then be recovered in the form of carbonate (LiCO) by adding gaseous CO.

[0013] To date, no efficient and economically viable technology is known that can recover Zn, Mn, Cd, Co, Ni, Li, and REEs from a mixture of used batteries, including alkaline, Zn-carbon, Ni-Cd, Ni-MH, Li-ion, and Li-M batteries, without expensive sorting steps. Canadian Patent Application No. 2,915,371 relates to a process for recovering Zn, Mn, Ni, and Cd from mixed waste batteries. The proposed process also integrates the recovery of lithium in the form of lithium carbonate and / or lithium phosphate downstream of the recovery of other metals. In addition, the proposed process includes the recovery of REE concentrate, which can be in the form of a mixed hydroxide and sulfate REE concentrate, an oxalate REE concentrate, or an oxide REE concentrate.

[0014] [Table 2] 1 H: Hydrometallurgy; P: Pyrometallurgy. Summary of the Invention

[0015] The differences between the process proposed herein and other known processes include: 1) the possibility of recovering almost all metals and rare earth elements of commercially valuable grades from battery waste; 2) very high metal recovery rates; and 3) operating conditions (e.g., reasonable chemical consumption, short retention times, and low temperatures) that are adapted to industrial processes compared to other existing options. Implementations of the process described herein are designed to recover metals and REEs of commercial interest in electronic waste. In particular, the process involves a selective leaching approach that first solubilizes REEs, followed by base metals, and finally precious metals. These metals can then be recovered by the techniques described herein for the treatment of battery waste.

[0016] According to an aspect, a process for treating electronic waste including at least one of used batteries and electronic waste is provided, the process including: fragmenting the electronic waste to reduce its size and producing fragmented waste including metal powders; recovering the metal powders from the fragmented waste, the metal powders including rare earth elements (REEs) and base metals; leaching the REEs and base metals from the metal powders to produce a residual metal-depleted solids and at least one of an REE-enriched leachate and a base metal-enriched leachate; extracting the REEs from the REE-enriched leachate to produce an REE-containing component and a secondary leachate; and selectively extracting the base metals from at least one of the base metal-enriched leachate and the secondary leachate to produce a plurality of extracted components, each extracted component including at least one of the base metals.

[0017] In some implementations, the basic metal comprises Cd, Co, Cu, Li, Mn, Ni, Zn, or any combination thereof.

[0018] In some implementations, the REEs include La, Nd, Sm, Ce, Pr, Y, Tb, Er, Sc, Gd, Eu, Dy, or any combination thereof.

[0019] In some implementations, fragmenting the e-waste includes at least one of shredding or shredding the e-waste.

[0020] In some implementations, recovering the metal powder includes separating the fragmented waste into coarse and fine fractions, the fine fractions including the metal powder.

[0021] In some implementations, separating the fragmented waste into coarse and fine fragments is performed by sieving or screening.

[0022] In some implementations, recovering the metal powder further includes recycling the coarse fragments as part of the e-waste to the fragmentation step b).

[0023] In some implementations, the fine fragments further include a low-density material comprising at least one of plastic and paper, and recovering the metal powder further includes separating the low-density material and the metal fragments from the fine fragments using a gravity-based method.

[0024] In some implementations, separating the low density material and metallic fragments from the fine fragments using a gravity-based method includes injecting air.

[0025] In some implementations, the metal fragments include an iron-based metal powder, and recovering the metal powder further includes separating the iron-based metal powder and the metal powder from the metal fragments by magnetic separation.

[0026] In some implementations, leaching the REEs and base metals from the metal powder includes simultaneously leaching the REEs and base metals by contacting the metal powder with a leach solution to solubilize at least a portion of both the REEs and base metals and produce a first solid-liquid mixture comprising an REE-enriched leachate and a metal-depleted residual solids, and separating the REE-enriched leachate and the metal-depleted residual solids by solid-liquid separation, wherein the REE-enriched leachate further comprises the base metal.

[0027] In some implementations, the leaching solution includes an inorganic acid.

[0028] In some implementations, the inorganic acid is selected from the group consisting of sulfuric acid (H2SO4), hydrochloric acid (HCl), and nitric acid (HNO3), preferably sulfuric acid.

[0029] In some implementations, the inorganic material has an acid concentration of 0.5N to 5N.

[0030] In some implementations, leaching the REEs includes using a leaching time of about 15 to 120 minutes, preferably about 30 minutes.

[0031] In some implementations, the ratio of the weight of the metal powder to the volume of the leach solution is between 50 g / L and 200 g / L.

[0032] In some implementations, the leaching solution further comprises hydrogen peroxide, sodium metabisulfite, or a combination thereof.

[0033] In some implementations, the electronic waste consists of used batteries.

[0034] In some implementations, leaching the REEs and base metals from the metal powder includes selectively leaching the REEs and base metals.

[0035] In some implementations, selectively leaching the REEs and the base metal includes leaching the REEs before leaching the base metal.

[0036] In some implementations, leaching the REEs includes contacting the metal powder with a first leach solution to solubilize at least a portion of the REEs, producing a first solid-liquid mixture including an REE-enriched leachate and an REE-depleted residual solids, and separating the REE-enriched leachate and the REE-depleted residual solids by solid-liquid separation.

[0037] In some implementations, the first leach solution includes an inorganic acid.

[0038] In some implementations, the inorganic acid is selected from the group consisting of sulfuric acid (H2SO4), hydrochloric acid (HCl), and nitric acid (HNO3).

[0039] In some implementations, the inorganic acid is a used or recycled acid.

[0040] In some implementations, the first leaching solution has an acid concentration of 0.2N to 1N.

[0041] In some implementations, leaching the REEs is carried out at a temperature of about 5-95°C, preferably at a temperature of about 20°C.

[0042] In some implementations, the ratio of the weight of the metal powder to the volume of the first leach solution is between 50 g / L and 200 g / L.

[0043] In some implementations, the ratio of the weight of the metal powder to the volume of the leach solution is 100 g / L.

[0044] In some implementations, leaching the base metals includes contacting the REE-depleted residual solids with a second leach solution to solubilize at least a portion of the base metals, producing a second solid-liquid mixture comprising the base metal-enriched leachate and the metal-depleted residual solids, and separating the base metal-enriched leachate and the metal-depleted residual solids by solid-liquid separation.

[0045] In some implementations, the second leach solution includes an inorganic acid and an oxidizing agent.

[0046] In some implementations, the oxidizing agent includes hydrogen peroxide.

[0047] In some implementations, the second leach solution includes an amount of oxidizing agent that is stoichiometrically sufficient to oxidize the base metal present in metallic form in the metal powder.

[0048] In some implementations, the second leach solution has an acid concentration of 1N to 5N.

[0049] In some implementations, the second leach solution has an acid concentration of 2N.

[0050] In some implementations, the second leach solution has a weight ratio of oxidizer to REE-depleted residual solids of 0.33 to 1.33.

[0051] In some implementations, the ratio of weight of REE-depleted residual solids to volume of second leach solution is between 50 g / L and 200 g / L.

[0052] In some implementations, the metal powder further comprises a precious metal, and the process further comprises leaching the precious metal from the metal-depleted residual solids to produce a precious-metal-enriched leachate and a metal concentrate.

[0053] In some implementations, leaching the precious metals includes contacting the metal-depleted residual solids with a third leach solution to solubilize at least a portion of the precious metals, producing a third solid-liquid mixture comprising a precious metal-enriched leachate and a metal concentrate, and separating the precious metal-enriched leachate and the metal concentrate by solid-liquid separation.

[0054] In some implementations, the third leaching solution has the same composition as the second leaching solution.

[0055] In some implementations, the third leach solution comprises a thiourea solution, a thiosulfate solution, a cyanide solution, or any mixture thereof.

[0056] In some implementations, the third leach solution has a leachant concentration of 0.10 to 0.50 grams of leachant per gram of metal-depleted residual solids.

[0057] In some implementations, the third leach solution further comprises an oxidizing agent in a stoichiometrically sufficient amount to oxidize the precious metal in metallic form.

[0058] In some implementations, the third leach solution has an oxidant concentration of at most 0.10 grams of oxidant per gram of metal-depleted residual solids.

[0059] In some implementations, the oxidizing agent is a ferric ion salt.

[0060] In some implementations, leaching of precious metals is carried out according to a solids content of the metal-depleted residual solids of 50 to 200 grams per liter of the third leach solution.

[0061] In some implementations, the third leach solution further comprises an inorganic acid.

[0062] In some implementations, the third leaching solution has an acid concentration of 0.1N to 0.5N.

[0063] In some implementations, the process further includes extracting precious metals from the precious metal-enriched leachate to produce a precious metal concentrate and a recyclable effluent.

[0064] In some implementations, the extraction of precious metals from the precious metal-enriched leachate includes using at least one of activated carbon adsorption, cementation, ion exchange, or electrodeposition.

[0065] In some implementations, the precious metal includes gold (Au), silver (Ag), platinum group metals (PGM), and any combination thereof.

[0066] In some implementations, the electronic waste includes used batteries and electronic waste.

[0067] In some implementations, extracting the REEs includes precipitating REE hydroxides and / or REE sulfates by adding a base solution to the REE-enriched leachate to produce a fourth solid-liquid mixture comprising the REE-containing component and a secondary leachate, wherein the REE-containing component comprises the REE hydroxides and / or REE sulfates, and separating the REE-containing component and the secondary leachate by solid-liquid separation.

[0068] In some implementations, extracting the REEs includes precipitating REE hydroxides and / or REE sulfates by adding a base solution to the metal-enriched leach solution to produce a fourth solid-liquid mixture; separating the REE hydroxides and / or REE sulfates from the fourth solid-liquid mixture via solid-liquid separation; re-dissolving the REE hydroxides and / or REE sulfates in an acid solution to form an REE-containing solution; adding oxalic acid and / or oxalate to the REE-containing solution to precipitate the REE oxalate to produce a fifth solid-liquid mixture containing the REE oxalate; and separating the REE oxalate as the REE-containing component from the fifth solid-liquid mixture via solid-liquid separation.

[0069] In some implementations, the base solution is selected from the group consisting of sodium hydroxide (NaOH), potassium hydroxide (KOH), calcium hydroxide (Ca(OH)2), and magnesium hydroxide (Mg(OH)2).

[0070] In some implementations, the base solution further comprises a sulfate salt.

[0071] In some implementations, extracting the REEs includes precipitating the REE oxalates by adding sulfate to the REE-enriched leachate to produce a fourth solid-liquid mixture containing the REE oxalates and a secondary leachate, and separating the REE oxalates and the secondary leachate as REE-containing components from the fourth solid-liquid mixture via solid-liquid separation.

[0072] In some implementations, the sulfate is sodium sulfate (Na2SO4), potassium sulfate (K2SO4), or any mixture thereof.

[0073] In some implementations, the process further includes calcining the REE-containing component to produce REE oxides.

[0074] In some implementations, the process further includes combining the base metal-enriched leachate with a secondary leachate to form a metal-enriched leachate, and selective extraction of the base metal is performed from the metal-enriched leachate.

[0075] In some implementations, selectively extracting the base metals includes sequentially extracting each one of the base metals to produce each extracted component containing one base metal.

[0076] In some implementations, the base metals include copper, zinc, cadmium, manganese, cobalt, nickel, and lithium, and the multiple extracted components sequentially produced by selective extraction of copper, zinc, cadmium, manganese, cobalt, nickel, and lithium include metallic copper (Cu), metallic zinc (Zn), metallic cadmium (Cd), cadmium sulfide concentrate (CdS), manganese carbonate concentrate (MnCO), cobalt oxide concentrate (CoO), nickel oxide concentrate (NiO), lithium carbonate (LiCO), and lithium phosphate (LiPO).

[0077] In some implementations, the e-waste includes used batteries, and the process further includes deactivating the used batteries before fragmenting the e-waste.

[0078] In some implementations, passivating the used battery includes placing the used battery in a saline or acid solution, freezing the used battery using liquid nitrogen, or a combination thereof.

[0079] In some implementations, leaching of REEs and base metals from metal powders involves performing multiple sequential acid leaching steps of the REEs and base metals.

[0080] In some implementations, leaching the REEs and base metals from the metal powder includes performing at least one acid leach of the REEs and base metals and at least one wash of the metal-depleted residual solids with water.

[0081] According to another aspect, a process for treating spent batteries containing basic metals and rare earth elements (REEs) includes the steps of: passivating the spent batteries; fragmenting the spent batteries to reduce their size and producing a fragmented waste material containing metal powders; recovering the metal powders from the fragmented waste material, the metal powders including the REEs and basic metals; and contacting the metal powders with a leach solution to solubilize at least a portion of both the REEs and the basic metals and produce a first solid-liquid mixture including a leachate and a metal-depleted residual solids content. the leachate and the metal-depleted residual solids by solid-liquid separation; extracting the REEs from the leachate to produce an REE-containing component and a secondary leachate, the secondary leachate comprising the base metals; and selectively extracting the base metals from the secondary leachate to produce a plurality of extracted components, each extracted component comprising at least one of the base metals.

[0082] In some implementations, passivating the used battery includes one of placing the used battery in a saline or acid solution or freezing the used battery using liquid nitrogen.

[0083] In some implementations, fragmentation of spent batteries is carried out in a non-oxidizing environment.

[0084] In some implementations, the non-oxidizing environment comprises a nitrogen atmosphere.

[0085] In some implementations, the process for the treatment of used batteries further includes at least one feature of some of the implementations described herein with respect to the treatment of electronic waste, including electronic waste.

[0086] In some implementations, the used battery is selected from the group consisting of an alkaline Zn / MnO2 battery, a Zn-C battery, a Ni-Cd battery, a Ni-MH battery, a Li-ion battery, a Li-M battery, a button cell, and any combination thereof.

[0087] In some implementations, the electronic waste comprises pieces of equipment selected from the group consisting of laptops, televisions, CRT monitors, LCD monitors, LED monitors, cell phones, smartphones, electronic tablets, solar photovoltaic (PV) panels, traditional hard drives (HDDs), solid-state drives (SSDs), and any combination thereof.

[0088] In some implementations, selectively extracting Cu includes adjusting the pH of the metal-enriched leachate by adding a base; adding a first organic solvent to create an aqueous-organic solution composed of an aqueous phase and an organic phase; extracting Cu from the aqueous phase to create a Cu-depleted aqueous solution and a Cu-enriched organic solution; separating the Cu-depleted aqueous solution from the Cu-enriched organic solution; stripping Cu from the Cu-enriched organic solution to form a Cu-depleted organic phase; and electrodepositing Cu onto a cathode to form an extractant comprising Cu, wherein the base is sodium hydroxide.

[0089] In some implementations, the first organic solvent is composed of 45% (v / v) mixed aldoxime-ketoxime reagent (e.g., LIX84-I) and 55% (v / v) kerosene.

[0090] In some implementations, extracting Cu from the aqueous phase comprises two stages of organic solvent extraction in which the ratio of the volume of the organic phase to the volume of the aqueous phase is 1.

[0091] In some implementations, separating the Cu-depleted aqueous solution from the Cu-rich organic solution is performed by decantation.

[0092] In some implementations, stripping Cu from the Cu-rich organic solution includes adding H2SO4 solution to a ratio of Cu-rich organic solution to H2SO4 solution of 1, obtaining a CuSO4 solution and a Cu-depleted organic phase, and separating the CuSO4 solution from the Cu-depleted organic phase.

[0093] In some implementations, electrodepositing Cu contained in a CuSO solution onto a cathode is performed using a Ti / IrO anode, a stainless steel cathode, and a 360 A / m 2 and producing effluent and electrodeposited Cu for recycling.

[0094] In some implementations, selectively extracting Zn includes adjusting the pH of the Cu-depleted aqueous solution by adding a base; adding a second organic solvent to create a second aqueous-organic solution composed of an aqueous phase and an organic phase; extracting Zn from the aqueous phase to create a Zn-depleted aqueous solution and a Zn-rich organic solution; separating the Zn-depleted aqueous solution from the Zn-rich organic solution; stripping Zn from the Zn-rich organic solution; and electrodepositing Zn onto a cathode.

[0095] In some implementations, the second organic solvent is composed of 20-30% (v / v) bis / 2,4,4-trimethylpentyl / phosphinic acid (Cyanex 272), 2% (v / v) tributyl phosphate (TBP), and 68-78% (v / v) kerosene.

[0096] In some implementations, extracting Zn from the aqueous phase comprises two stages of organic solvent extraction, with the volume of the organic phase to the volume of the aqueous phase being 2.

[0097] In some implementations, separating the Zn-depleted aqueous solution from the Zn-rich organic solution is performed by decantation.

[0098] In some implementations, stripping Zn from the Zn-rich organic solution includes adding H2SO4 solution such that the ratio of the volume of the Zn-rich organic solution to the volume of the H2SO4 solution is 2, obtaining a ZnSO4 solution and a Zn-depleted organic phase, stripping residual iron from the Zn-depleted organic phase to form a resultant iron solution, and separating the solution from the Zn-depleted organic phase.

[0099] In some implementations, stripping the residual iron includes forming an FeSO4 solution.

[0100] In some implementations, separating the iron solution from the Zn-depleted organic phase is performed by decantation.

[0101] In some implementations, electrodepositing Zn onto a cathode is performed using a Ti / IrO2 anode, a stainless steel or aluminum cathode, and a current of 250-750 A / m 2 and producing effluent and electrodeposited Zn for recycling.

[0102] In some implementations, selectively extracting Cd and Mn includes adjusting the pH of the Zn-depleted aqueous solution by adding a base; adding a third organic solvent to create a third aqueous-organic solution composed of an aqueous phase and an organic phase; extracting Cd and Mn from the aqueous phase to create a Cd-Mn-depleted aqueous solution and a Cd-Mn-enriched organic solution; separating the Cd-Mn-depleted aqueous solution from the Cd-Mn-enriched organic solution; scrubbing the Cd-Mn-enriched organic solution; stripping Cd and Mn from the Cd-Mn-enriched organic solution; electrodepositing Cd onto a cathode; separating the Mn solution; and treating the Mn solution.

[0103] In some implementations, the third organic solvent is composed of 30% (v / v) di-(2-ethylhexyl)phosphoric acid (D2EHPA), 5% (v / v) TBP, and 65% (v / v) kerosene.

[0104] In some implementations, extracting Cd and Mn from the aqueous phase comprises two stages of organic solvent extraction, with the volume of the organic phase to the volume of the aqueous phase being 2.

[0105] In some implementations, separating the Cd—Mn-depleted aqueous solution from the Cd—Mn-rich organic solution is performed by decantation.

[0106] In some implementations, scrubbing the Cd—Mn-rich organic solution includes adding H2SO4 solution in a ratio of a volume of the Cd—Mn-rich organic solution to a volume of the H2SO4 solution of 20, and removing impurities.

[0107] In some implementations, stripping Cd and Mn from the Cd—Mn-rich organic solution includes adding H2SO4 solution to the Cd—Mn-rich organic solution such that the ratio of the volume of the Cd—Mn-rich organic solution to the volume of the H2SO4 solution is 4, and obtaining a Cd—Mn sulfate solution.

[0108] In some implementations, electrodepositing Cd involves using a Ti / IrO2 anode, a stainless steel or aluminum cathode, and a current of 360-370 A / m 2 and producing an electrodeposited Cd and Mn solution.

[0109] In some implementations, stripping Cd and Mn from the Cd—Mn-rich organic solution includes precipitating Cd from the Cd—Mn-rich organic solution as a cadmium sulfide (CdS) concentrate by adding a first base and a sulfide salt to the Cd—Mn-rich organic solution, and separating the CdS from the waste stream.

[0110] In some implementations, separating the CdS concentrate is performed by using solid-liquid separation.

[0111] In some implementations, the sulfide salt is NaHS or Na2S.

[0112] In some implementations, the Mn solution is a MnSO4 solution.

[0113] In some implementations, treating the Mn solution includes precipitating MnCO by adding a first base and a carbonate salt, and separating the MnCO from a waste stream.

[0114] In some implementations, the first base is selected from the group consisting of sodium hydroxide (NaOH), potassium hydroxide (KOH), calcium hydroxide (Ca(OH)), and magnesium hydroxide (Mg(OH)).

[0115] In some implementations, separating the MnCO is performed by using solid-liquid separation.

[0116] In some implementations, the carbonate is Na2CO3 or K2CO3.

[0117] In some implementations, selectively extracting Co and Ni includes adjusting the pH of the Cd—Mn-depleted aqueous solution by adding a base; adding a fourth organic solvent to create a fourth aqueous-organic solution composed of an aqueous phase and an organic phase; extracting Co and Ni from the aqueous phase to create a Co—Ni-depleted aqueous solution and a Co—Ni-rich organic solution; separating the Cd—Mn-depleted aqueous solution from the Cd—Mn-rich organic solution; scrubbing the Co—Ni-rich organic solution to remove Ni from the Co—Ni-rich organic solution to create a Co-rich organic solution and a Co-depleted aqueous solution; stripping Co from the Co-rich organic solution; precipitating Co by adding oxalate or oxalic acid to form a Co-oxalate precipitate; separating the Co-oxalate precipitate to form a Co-oxalate concentrate; calcining the Co-oxalate concentrate to obtain Co oxide (CoO); and treating the Co-depleted aqueous solution.

[0118] In some implementations, the fourth organic solvent is composed of 10% (v / v) bis / 2,4,4-trimethylpentyl / phosphinic acid (Cyanex 272), 2% (v / v) TBP, and 88% (v / v) kerosene.

[0119] In some implementations, extracting Co and Ni from the aqueous phase comprises two stages of organic solvent extraction with a volume of organic phase to volume of aqueous phase of 0.5.

[0120] In some implementations, separating the Co—Ni depleted aqueous solution from the Cd—Mn rich organic solution is performed by decantation.

[0121] In some implementations, scrubbing the Co—Ni-rich organic solution includes adding CoSO to obtain a ratio of the volume of the Co—Ni-rich organic solution to the volume of CoSO of 4, and separating the Co-rich organic solution from the Co-depleted aqueous solution.

[0122] In some implementations, stripping Co from the Co-rich organic solution includes adding H2SO4 solution such that the ratio of the volume of the Co-rich organic solution to the volume of the H2SO4 solution is 2 to form a CoSO4 solution.

[0123] In some implementations, precipitating Co includes adding oxalate or oxalic acid to the CoSO solution to form a Co oxalate precipitate.

[0124] In some implementations, separating the Co-oxalate precipitate is performed using solid-liquid separation.

[0125] In some implementations, treating the Co-depleted aqueous solution includes precipitating Ni oxalate by adding oxalate or oxalic acid, separating the Ni oxalate from the Co-Ni depleted solution by using solid-liquid separation, and calcining the Ni oxalate to obtain Ni oxide (NiO).

[0126] In some implementations, selectively extracting LiCO and LiPO includes: Precipitating metal hydroxides in the Co-Ni-reduced solution by adding a second base solution, separating the metal hydroxides from the resulting lithium sulfate solution, adjusting the pH to a pH of 9-10 using a second base and precipitating LiCO3 by adding carbonate or a concentrated solution of carbonate, separating the LiCO3 from the LiCO3-reduced solution, and precipitating Li3PO4 by adding phosphate to the LiCO3-reduced solution, and separating the Li3PO4 from the lithium-reduced solution.

[0127] In some implementations, the second base solution is selected from the group consisting of sodium hydroxide (NaOH), potassium hydroxide (KOH), calcium hydroxide (Ca(OH)), and magnesium hydroxide (Mg(OH)).

[0128] In some implementations, separating the metal hydroxide from the resulting lithium sulfate solution is performed using solid-liquid separation.

[0129] In some implementations, the carbonate is sodium carbonate (Na2CO3) or potassium carbonate (K2CO3).

[0130] In some implementations, separating LiCO3 and separating Li3PO4 from the lithium-depleted solution is performed using solid-liquid separation.

[0131] In some implementations, the phosphate is sodium phosphate (Na3PO4).

[0132] In some implementations, extracting Li3PO4 is performed without extracting LiCO3.

[0133] In some implementations, the residual metal-depleted solids is a magnesium dioxide (MnO2) concentrate.

[0134] In some implementations, the REE-containing component is an REE oxide concentrate.

[0135] In some implementations, fragmentation involves crushing and / or chopping the fine pieces until a particle size of less than 5 mm is obtained.

[0136] In some implementations, recovering the metal powder further includes a water-based milling process.

[0137] In some implementations, the milling step is carried out by vigorously mixing a solution containing 30-70% fine fragments in water for about 2-15 minutes.

[0138] In some implementations, the solid-liquid separation is selected from the group consisting of filtration, centrifugation, and sedimentation techniques.

[0139] In some implementations, the metal-depleted residual solids are treated by sequential leaching steps with acid or washed with water.

[0140] In some implementations, leaching the REEs and base metals includes reducing the volume of the REE-enriched leachate using a countercurrent washing process or a countercurrent leaching process.

[0141] In some implementations, the process further includes washing the fine fraction before leaching.

[0142] In some implementations, leaching of REEs and base metals is carried out at a temperature of about 5-95°C, preferably at a temperature of about 20°C, for solubilization of REEs.

[0143] In some implementations, leaching of the REEs and base metals includes using a leaching time of about 15 to 120 minutes, preferably about 30 minutes, for solubilization of the REEs.

[0144] In some implementations, leaching of the REEs and base metals includes using a leaching time of about 60 to 360 minutes, preferably about 180 minutes, for solubilization of the base metals.

[0145] In some implementations, leaching of precious metals involves using a temperature of 5 to 95°C, preferably a temperature of about 20°C, for solubilization of precious metals.

[0146] In some implementations, leaching of precious metals involves using a thiourea concentration of about 0.10 to 0.50 g / g of metal-depleted residual solids, preferably about 0.25 g / g of metal-depleted residual solids, for solubilization of precious metals.

[0147] In some implementations, leaching of the precious metals includes using a leaching time of about 30 to 240 minutes, preferably about 120 minutes, for solubilization of the precious metals.

[0148] In some implementations, leaching of precious metals involves proceeding at a solids concentration of about 50-200 g of metal-depleted residual solids per liter of solution, preferably about 130 g of metal-depleted residual solids per liter of solution, for solubilization of the precious metals.

[0149] In some implementations, precipitating the REE hydroxides and / or REE sulfates is carried out at a pH of at least 1.2, preferably at a pH of about 1.5.

[0150] In some implementations, precipitating the REE hydroxides and / or REE sulfates is carried out at a temperature of 5 to 95°C, preferably at a temperature of about 20°C.

[0151] In some implementations, precipitating the REE hydroxides and / or REE sulfates is carried out using a precipitation time of about 5 to 60 minutes, preferably about 10 minutes.

[0152] In some implementations, redissolving the REE hydroxides and / or REE sulfates includes using a hydrochloric acid solution.

[0153] In some implementations, precipitating the REE oxalates is carried out at a pH of about 0.75.

[0154] In some implementations, precipitating the REE oxalates is carried out at a temperature of about 5-95°C, preferably at a temperature of about 20°C.

[0155] In some implementations, precipitating the REE oxalates includes using a precipitation time of about 5 to 60 minutes, preferably about 10 minutes.

[0156] In some implementations, precipitating the REE oxalates is performed without precipitating the REE hydroxides and REE sulfates.

[0157] In some implementations, extracting Cu from the aqueous phase using the first organic solvent is carried out at a pH of about 2.

[0158] In some implementations, stripping Cu from a Cu-rich organic solution includes using H2SO4 solutions at concentrations of 0.5N and 5N.

[0159] In some implementations, extracting Cu from the aqueous phase is carried out at a temperature of about 20°C to 60°C, preferably at a temperature of about 20°C.

[0160] In some implementations, stripping Cu from the Cu-rich organic solution includes using a contact time of about 10 minutes.

[0161] In some implementations, electrodepositing Cu onto the cathode is carried out at a pH of about 2 and a temperature of about 20°C to 60°C, preferably at a temperature of about 20°C.

[0162] In some implementations, electrodepositing Cu onto the cathode is carried out using a contact time of about 120 minutes.

[0163] In some implementations, extracting Zn from the aqueous phase is carried out at a pH of about 2.0 to 2.5.

[0164] In some implementations, extracting Zn from the aqueous phase is carried out at a temperature of about 40°C to 60°C, preferably at a temperature of about 50°C.

[0165] In some implementations, stripping Zn from the Zn-rich organic solution includes using a H2SO4 solution at a concentration of 2N.

[0166] In some implementations, stripping Fe from the Zn-rich organic solution includes using a H2SO4 solution at a concentration of 2N and a ratio of the volume of the Zn-rich organic solution to the volume of the H2SO4 solution of 2.

[0167] In some implementations, stripping Fe from the Zn-rich organic solution includes using a contact time of about 10 minutes.

[0168] In some implementations, electrodepositing Zn is carried out at a pH of about 2.0 and a temperature of about 40°C to 60°C, preferably about 50°C.

[0169] In some implementations, electrodepositing Zn includes using a contact time of about 120 to 180 minutes.

[0170] In some implementations, extracting Cd and Mn from the aqueous phase is carried out at a pH of about 2.7-2.9 and a temperature of about 40°C-60°C, preferably about 50°C.

[0171] In some implementations, scrubbing the Cd—Mn-rich organic solution is carried out at an equilibrium pH of about 2.3.

[0172] In some implementations, stripping Cd and Mn from the Cd-Mn rich organic solution includes using a H2SO4 solution at a concentration of about 0.5-0.6N.

[0173] In some implementations, stripping Cd and Mn from the Cd—Mn-rich organic solution includes using a contact time of about 10 minutes.

[0174] In some implementations, electrodepositing Cd onto the cathode is carried out at a pH of about 2.0 and a temperature of about 40°C to 60°C, preferably about 50°C.

[0175] In some implementations, electrodepositing Cd onto the cathode is carried out using a contact time of about 120 to 240 minutes.

[0176] In some implementations, precipitating Mn is carried out at a pH of about 8-9 and at room temperature.

[0177] In some implementations, precipitating Mn includes using a retention time of about 15 to 240 minutes.

[0178] In some implementations, extracting Co and Ni from the aqueous phase is carried out at a pH of about 5.2-5.5 and a temperature of about 40°C-60°C, preferably about 50°C.

[0179] In some implementations, stripping Co from the Co-rich organic solution includes using a H2SO4 solution at a concentration of 0.25N.

[0180] In some implementations, stripping Co from the Co-rich organic solution includes using a contact time of 10 minutes.

[0181] In some implementations, precipitating Co to form a Co oxalate precipitate includes using a retention time of about 15 to 240 minutes.

[0182] In some implementations, precipitating Co to form a Co oxalate precipitate is carried out at room temperature.

[0183] In some implementations, precipitating Ni oxalate includes using a retention time of about 15 to 240 minutes.

[0184] In some implementations, precipitating Ni oxalate is carried out at room temperature.

[0185] In some implementations, precipitating metal hydroxides in the reduced Co—Ni solution is carried out at a pH of about 7-10.

[0186] In some implementations, precipitating LiCO3 is carried out at a temperature of about 0°C to 100°C, preferably at a temperature of about 50°C.

[0187] In some implementations, precipitating the LiCO includes using a hold time of about 45 minutes.

[0188] While the present technology will be described in conjunction with exemplary embodiments, it will be understood that it is not intended to limit the scope of the present invention to such embodiments. On the contrary, it is intended to cover all alternatives, modifications, and equivalents that may be included as defined by this description. Objects, advantages, and other features of the present technology will become more apparent and be better understood upon reading the following non-limiting description of the invention, which is given with reference to the accompanying drawings. [Brief explanation of the drawings]

[0189] The accompanying figures illustrate various features, aspects and implementations of the technology described herein. [Figure 1] FIG. 1 is a schematic process flow diagram of a general process for fragmenting e-waste into fragmented waste and recovering metal powder therefrom. [Figure 2] FIG. 1 is a schematic process flow diagram of a general step for leaching both rare earth elements and base metals from a metal powder, thereby producing an REE-enriched leachate containing base metals and rare earth elements. [Figure 3] FIG. 1 is a schematic process flow diagram of the general steps for selectively leaching rare earth elements, base metals, and precious metals from metal powders, thereby producing an REE-enriched leachate, a base metal-enriched leachate, and a precious metal-enriched leachate. [Figure 4] FIG. 1 is a schematic process flow diagram of the general steps for extracting rare earth elements from an REE-enriched leachate, thereby producing a secondary leachate. [Figure 5] FIG. 1 is a schematic process flow diagram of a general step for selectively extracting base metals from at least one of a base-metal-enriched leachate and a secondary leachate to recover multiple extracted components. [Figure 6] FIG. 1 is a schematic process flow diagram of a general step for extracting metallic copper from at least one of a secondary leachate and a basic metal-enriched leachate to produce a Cu-depleted solution. [Figure 7] FIG. 1 is a schematic process flow diagram of the general steps for extracting zinc metal from a Cu-depleted solution to produce a Zn-depleted solution. [Figure 8] FIG. 1 is a schematic process flow diagram of the general steps for extracting cadmium metal from a Zn-depleted solution and producing a MnSO solution. [Figure 9] FIG. 1 is a schematic process flow diagram of the general steps for extracting manganese as the MnCO component from a MnSO solution to produce a Mn-depleted solution. [Figure 10] FIG. 1 is a schematic process flow diagram of a general step for extracting cobalt as the CoO component from a Mn-depleted solution to produce a Co-depleted solution. [Figure 11] FIG. 1 is a schematic process flow diagram of the general steps for extracting nickel as the NiO component from a Co-depleted solution and producing a tertiary leach solution. [Figure 12]FIG. 1 is a schematic process flow diagram of the general steps for extracting lithium in the form of lithium carbonate (LiCO) and / or lithium phosphate (LiPO) from the tertiary leach solution and producing an effluent for recycling. [Figure 13] FIG. 1 is a schematic process flow diagram of the general steps for extracting precious metals from a precious metal leach solution in the form of a precious metal concentrate and producing an effluent for recycling. DETAILED DESCRIPTION OF THE INVENTION

[0190] The present technology provides an e-waste processing solution that allows for the combination of different sources of e-waste without the need for waste sorting. For example, the e-waste can be a mixture of unsorted used batteries, various used electronic components, or a combination thereof. The present technology further allows for the tailoring of the process according to the nature of the e-waste to selectively extract and recover at least both base metals and rare earth elements (RREs). Precious metals, if present, can also be subsequently extracted and recovered. The base metals, RREs, and precious metals are recovered as multiple extractable components, which can be the elements in metallic form (e.g., metallic copper) or can include the elements in complexed, oxidized, or carbonated form (e.g., NiO, CoCO). Note that the extractable components may be directly priced for use in secondary processes, thereby reducing or avoiding the disposal of hazardous metals in landfills.

[0191] In some implementations, the used batteries may be selected from the group consisting of alkaline Zn / MnO2 batteries, Zn-C batteries, Ni-Cd batteries, Ni-MH batteries, Li-ion batteries, Li-M batteries, and any combination thereof. The electronic waste may include pieces of equipment selected from the group consisting of laptops, televisions, CRT monitors, LCD monitors, LED monitors, mobile phones, smartphones, electronic tablets, photovoltaic (PV) panels, traditional hard drives (HDDs), solid-state drives (SSDs), and any combination thereof.

[0192] The general steps of the process proposed herein include the production of metal powders from e-waste, selective leaching of RRE, base metals, and precious metals (if present) to recover a multi-metal-enriched leachate, and then selective extraction of the RRE, base metals, and precious metals (if present) in the form of multiple extracted components.

[0193] Metal powder production and recovery As seen in Figure 1, the production of metal powder involves fragmenting (4) electronic waste (2) to reduce its size and produce fragmented waste (6) containing metal powder. The metal powder may be recovered in the fine fraction (10) resulting from separating (8) the fragmented waste (6) into coarse fraction (12) and fine fraction (10). It is further noted that the fine fraction (8) of the fragmented waste (6) may further include low-density materials and ferrous metals.

[0194] The low-density material may include at least one of plastic and paper. The ferrous metals in the fine fragment (10) may be present as ferrous scrap. The low-density material and ferrous metals may be separated from the fine fragment (10) according to techniques described further herein or any other conventional separation techniques known in the art to separately recover the metal powder.

[0195] The metal powder recovered from the fine pieces (10) of the fragmented waste (6) includes basic metals, REEs, and optionally precious metals. More specifically, when the e-waste (2) consists of or includes a mixture of used batteries, including alkaline batteries, Zn-C batteries, Ni-Cd batteries, Ni-MH batteries, Li-ion batteries, Li-M batteries, or any combination thereof, the metal powder may include basic metals and REEs. References to basic metals herein include Al, Ba, Cd, Co, Cu, Fe, K, Li, Mg, Mn, Na, Ni, S, Zn, or any combination thereof. References to "basic metals" should be understood to refer to one or more metals in this category. References to REEs herein include La, Nd, Sm, Ce, Pr, Y, Tb, Er, Sc, Gd, Eu, Dy, or any combination thereof. References to "REEs" should be understood to refer to one or more metals in this category. If the e-waste further includes used electronic components, the metal powder may further include a precious metal. Noble metals referred to herein include Ag, Au, Pd, Pt, or any combination thereof. It should be understood that reference to a "noble metal" refers to one or more metal compounds of this category.

[0196] As shown in Figure 1, if the e-waste (2) to be processed includes both unsorted used batteries (2a) and electronic waste (2b), the metal powder (14) may be referred to as battery powder and electronics powder (14a). As shown in Figure 2, if the e-waste to be processed consists of unsorted used batteries, the metal powder (14) may be referred to as fine battery powder (14b). As shown in Figure 3, if the e-waste to be processed consists of electronic waste, the metal powder (14) may be referred to as electronics powder (14c).

[0197] In some implementations, the e-waste includes unsorted used batteries, and the process may further include deactivating the used batteries before fragmenting the e-waste. In implementations where the e-waste (2) includes used batteries (2a), as seen in FIG. 1 , deactivation (16) of the used batteries (2a) may be performed before fragmenting (4) the e-waste (2) to reduce the risk of explosion and fire during fragmentation (4). This risk is primarily associated with the presence of lithium and Ni-MH batteries in the used batteries.

[0198] For example, a mixture of used batteries can be subjected to a deactivation process by freezing the used batteries, e.g., using liquid nitrogen, and then gently thawing the used batteries. In another example, deactivation can include discharging the used batteries by immersing them in a saline or acid solution for several hours. The saline or acid solution can be reused for some battery deactivation processes. In yet another example, deactivation and fragmentation can be performed simultaneously by fragmenting the used batteries in a non-oxidizing environment, e.g., a nitrogen atmosphere.

[0199] Fragmenting e-waste into fragmented waste can be performed according to various techniques available in the art. Fragmentation is performed to reduce the size of the e-waste and generate fragments of various sizes. The fragments can be generated by at least one of shredding, shredding, cutting, and grinding. Referring to FIG. 1 , e-waste (2b) can be combined with deactivated batteries (2c) for further fragmentation (4) by shredding and shredding without the need for a deactivation step to generate fragmented waste (6). The fragmented waste (6) includes coarse fragments (12) and fine fragments (10). The fine fragments (10) include metal powder (14) and other components, including plastic, paper, and metal fragments, including ferrous metal powders.

[0200] Still referring to FIG. 1, the fragmentation step (4) is followed by multiple steps to recover the metal powder (14). Depending on the nature of the fragmented waste (6), recovery can include separating (8) fine fractions (10) from the fragmented waste (6) and removing unwanted fragments from the fine fractions (10) until the metal powder (14) is recovered. Recovery can include separating (8) the fragmented waste (6) into coarse fractions (12) and fine fractions (10). Separation (8) can be performed according to various techniques available in the art to recover fine fractions having an average particle size of less than 1 mm. For example, separation (8) can include sieving. This separation (8) step can use sieves with openings of less than 5 mm, preferably 1-2 mm. The coarse fractions (12) can be recycled back to the fragmentation step (4) to obtain a suitable particle size.

[0201] In some implementations, as shown in FIG. 1 , recovering the metal powder (14) may further include subjecting the fine fraction (10) to a grinding step (18) using water (20) to remove and separate debris adhering to the surfaces of the particles of the fine fraction (10) and / or to break down agglomerates of the particles of the fine fraction (10). The grinding step (18) may be carried out by preparing a solution of the fine fraction (10) and water (20) having a solids content ranging from 30% to 70% to form a fine fraction solution. This fine fraction solution may then be vigorously stirred in a stirred tank reactor for 2 to 15 minutes.

[0202] 1, after grinding is complete, the fine fragment solution (22) can be subjected to gravitational separation (24), such as a flotation separation process, to separate the fine fragment solution (22) into low-density materials (26) and resulting metal fragments (28). Gravimetric separation (24) can result from any gravity-based method and can involve simply allowing the solution (22) from the grinding process (18) to settle and separating the floating low-density materials (26), which may include plastic, paper, and cardboard particles remaining on top of the solution (22). Optionally, separation of the low-density materials (26) can be further facilitated by injecting air into the fine fragment solution (22).

[0203] Still referring to FIG. 1 , recovery of the metal powder (14) may further include magnetic separation (30) to separate the metal fragments (28) into ferrous metal powder (32) (including scrap of ferrous metals, such as iron scrap) and metal powder contained in non-magnetic metal fragments (34). Note that when grinding is used, the metal fragments (34) are provided in a solution containing non-ferrous metals. Therefore, the metal powder (14) may be further recovered via solid-liquid separation (36), such as filtration, centrifugation, or sedimentation. Optionally, water (38) used in the grinding, gravity-based, and magnetic separation steps may be recovered. Such used water (38) may contain dissolved solids (42), which may be partially removed by wastewater treatment (40), including precipitation, and the treated water (44) may be recycled to other processing loops.

[0204] In another embodiment, used batteries and electronic waste can be processed separately to recover the metal powder.

[0205] Leaching of base metals and REEs The process for treating e-waste further includes leaching REEs and base metals from the metal powder to produce a residual metal-reduced solids fraction and at least one of an REE-enriched leachate and a base metal-enriched leachate. When the e-waste includes electronic waste, precious metals are further leached from the residual metal-reduced solids fraction. In embodiments where the e-waste includes or consists of used batteries, the leaching of REEs from the metal powder and the leaching of base metals can be performed simultaneously. In embodiments where the e-waste includes or consists of electronic waste, the leaching of REEs, the leaching of base metals, and the leaching of precious metals can be performed selectively, more particularly sequentially.

[0206] Leaching refers to the use of a leachate to solubilize metals contained in solid particles of metal powder derived from e-waste encompassed herein. In some implementations, the leachate solution for solubilizing REEs can be the same as the leachate solution for solubilizing base metals, thereby forming a leachate containing solubilized REEs and base metals. The leachate solution for solubilizing REEs and / or base metals can be a dilute sulfuric acid solution. The leachate solution for solubilizing precious metals contained in metal powders can be a dilute thiourea solution. Acid leaching, as used herein, refers to the use of an inorganic acid as at least part of the acid leach solution. When referring to multiple acid leach steps to perform leaching of REEs and base metals, the nature of the inorganic acid, the composition, and the concentration of the acid leach solution can vary from one acid leach step to the next.

[0207] In some implementations, leaching of metals (REEs, base metals, and / or precious metals) or washing of residual solids can be carried out using a countercurrent process. For example, metals from e-waste (or residual solids from a previous leaching / washing / extraction step) can be leached by contacting them with a leach solution that flows countercurrently to the metal powder. Unless otherwise indicated, leaching can further include washing the residual solids with water after contacting them with the leach solution.

[0208] In one embodiment, leaching the base metal and REE from the metal powder may include contacting the metal powder (e.g., fine battery powder) with a leach solution to solubilize at least a portion of both the REE and the base metal, thereby simultaneously leaching the REE and the base metal. For example, with reference to Figure 2, fine battery powder (14b) may be contacted with an acid solution (46) for acid leaching (48) to produce a first solid-liquid mixture (50) comprising an REE-enriched leachate (52) and a metal-depleted residual solids (54).

[0209] Leaching (48) of the base metals and REEs from the metal powder (14) can be carried out by mixing a predetermined amount of the metal powder (14) with an inorganic acid, such as sulfuric acid, hydrochloric acid, or nitric acid. For example, sulfuric acid can be used. The leaching solution (46) can have an acid concentration of 0.5 to 5.0 N. Leaching of the base metals and REEs can be carried out with a ratio of metal powder weight to leach solution volume of 50 g / L to 200 g / L and can involve mixing for 15 to 120 minutes. Preferably, a 4.0 N sulfuric acid concentration is used with a solids content of approximately 100 g / L and a reaction time of 30 minutes. The temperature can be maintained between 5 and 95°C during leaching of the base metals and REEs, with a recommended temperature being 20°C.

[0210] In some implementations, the leaching solution for leaching base metals and REEs from metal powders further comprises hydrogen peroxide, sodium metabisulfite, or a combination thereof in addition to an inorganic acid.

[0211] After contacting the metal powder (14) with the leach solution (46), and still referring to FIG. 2, the first solid-liquid mixture (50) can be separated into a metal-depleted residual solids fraction (54) and an REE-enriched leachate (52) by solid-liquid separation (56), where the REE-enriched leachate (52) contains both REEs and base metals. The solid-liquid separation (56) described herein can be carried out according to techniques known in the art, such as, for example, filtration, centrifugation, and / or sedimentation. The metal-depleted residual solids fraction (54) can be a MnO2 concentrate.

[0212] In some implementations, and still referring to Figure 2, the metal-depleted residual solids (54) may undergo an additional rinse (58), for example with water (60), and further solid-liquid separation (62) to recover a cleaner metal-depleted residual solids (64) (e.g., MnO concentrate). The wash effluent produced as a flowable fraction (66) may be recycled to the acid leach (48).

[0213] It should be noted that the metal-depleted residual solids may contain residual REEs and / or base metals, and therefore leaching may involve multiple sequential leaching and / or washing steps to maximize the amount of REEs and base metals solubilized in the REE-enriched leachate. For example, the metal-depleted residual solids may be subjected to further sequential leaching steps (e.g., 0 to 4 additional steps) with acid and / or rinsed by one or more mixing steps with water. The volume of solution generated may be reduced using the principle of countercurrent washing and leaching steps. Optionally, the process may include washing the metal powder in water prior to acid leaching to remove a significant proportion of the potassium and sodium present in the metal powder.

[0214] In another embodiment, more particularly when the e-waste includes electronic waste, leaching of REEs, base metals, and precious metals can be performed by a selective leaching process. For example, selectively leaching REEs and base metals can include leaching the REEs before leaching the base metals. Referring to FIG. 3 , where the metal powder (14) is electronics powder (14c), the REE teachings can include contacting the metal powder (14) with a first leach solution (68), e.g., an acid solution, to perform a first leach (70) to solubilize at least a portion of the REEs, thereby producing a first solid-liquid mixture (72) including an REE-enriched leachate (74) and an REE-depleted residual solids (76). The REE-enriched leachate (74) contains at least a portion of the REEs from the metal powder. The REE-enriched leachate (74) and the REE-depleted residual solids (76) can then be separated by a first solid-liquid separation (78).

[0215] The first leach solution (68) may comprise an inorganic acid solution (having an acid concentration of 0.2-1.0N). Leaching of the REEs may be carried out using a ratio of the weight of the metal powder to the volume of the first leach solution of 50 g / L to 200 g / L. Leaching of the REEs may also include a mixing step lasting 15-120 minutes. Preferably, a sulfuric acid concentration of 0.4N is used with a solids content of approximately 100 g / L and a reaction time of 30 minutes. The temperature of the mixture may be maintained between 5 and 95°C during REE leaching, with a recommended temperature being 20°C. After leaching of the REEs, the REE-depleted residual solids and the REE-enriched leachate may be separated by solid-liquid separation.

[0216] 3, the second step of selectively leaching the REEs and base metals may comprise leaching the base metals (80) and may include contacting the REE-depleted residual solids (76) with a second leach solution (82) to solubilize at least a portion of the base metals and produce a second solid-liquid mixture (84) comprising a base metal-enriched leachate (86) and a metal-depleted residual solids (88). The base metal-enriched leachate (86) and the metal-depleted residual solids (88) may then be separated by another second solid-liquid separation (90).

[0217] In some implementations, the second leach solution (82) may include an inorganic acid solution and an oxidizing agent. The oxidizing agent may be hydrogen peroxide. The inorganic acid solution may be a sulfuric acid, hydrochloric acid, or nitric acid solution. Leaching of basic metals may require an acid concentration of the second leach solution between 1.0 N and 5.0 N, and a weight ratio of oxidizing agent to REE-depleted residual solids between 0.33 and 1.33.

[0218] For example, assuming the basic metal content is equal to the copper content, the amount of oxidizer to add can be determined based on stoichiometry by knowing the basic metal content present in the powder mixture as follows: Cu(s)+H2O2(aq)+H2SO4(aq)→CuSO4(aq)+2H2O(aq) (1)

[0219] The amount of oxidizing agent added to the second leach solution may be stoichiometrically sufficient to oxidize the basic metals present in metallic form in the metal powder, and more precisely in the REE-depleted residual solids.

[0220] In some implementations, a ratio of the weight of the REE-depleted residual solids to the volume of the second leach solution of 50 g / L to 200 g / L may be used for leaching the base metals and a mixing time of 60 to 360 minutes. Preferably, sulfuric acid with an acid concentration of 2.0 N and 0.67 g H2O2 / g per gram of REE-depleted residual solids may be used with a solids content of approximately 100 g / L and a reaction time of 180 minutes. The temperature may be maintained between 5 and 95°C during leaching of the base metals, with a recommended temperature being 80°C. After leaching of the base metals, the metal-depleted residual solids and the base-metal-enriched leach solution may be separated by solid-liquid separation.

[0221] In some embodiments, leaching the REEs and base metals from the metal powder may comprise performing multiple sequential acid leaching steps of the REEs and base metals, hi other embodiments, leaching the REEs and base metals from the metal powder may comprise performing at least one acid leach of the REEs and base metals and at least one wash of the metal-depleted residual solids with water.

[0222] In some embodiments, the metal powder may include a precious metal. Accordingly, the third leaching step may comprise leaching the precious metal from the metal-depleted residual solids to produce a precious-metal-enriched leachate and a metal concentrate. Still referring to Figure 3, for example, leaching the precious metal (92) may comprise contacting the metal-depleted residual solids (88) with a third leach solution (94) to solubilize at least a portion of the precious metal and produce a third solid-liquid mixture (96) comprising a precious-metal-enriched leachate (98) and a metal concentrate (100).

[0223] The third leach solution may comprise or consist of a thiourea solution, a thiosulfate solution, a cyanide solution, or any mixture thereof. Leaching of precious metals may be achieved by forming a soluble complex with the thiourea solution, the thiosulfate solution, or the cyanide solution. The use of a thiourea solution may be recommended for the extraction of precious metals, as the extraction can be carried out in an acidic environment, avoiding a significant increase in pH.

[0224] In some implementations, the third leach solution may further include an oxidizing agent in a stoichiometrically sufficient amount to oxidize the precious metal in metallic form.

[0225] Thiourea can also be used for metal leaching due to its low toxicity (LD50 (thiourea) = 125 mg / kg and LD50 (potassium cyanide) = 5 mg / kg), promising recovery, faster reaction rate, and high selectivity.

[0226] The precious metals may include gold (Au), silver (Ag), and platinum group metals (PGM), or any combination thereof.

[0227] Leaching of precious metals can be carried out using iron as the oxidant according to the following scheme: Au(s) + 2CS(NH2)2 + Fe 3+ →Au(CS(NH2)2) 2+ +Fe 2+ (2) Ag(s)+3CS(NH2)2+Fe 3+ →Ag(CS(NH2)2) 3+ +Fe 2+ (3)

[0228] Noble metal leaching can also be carried out in the presence of cyanide according to the following scheme: 4Au(s)+8CN - +O2+2H2O→4Au(CN) 2- +4OH - (4) 4Ag(s)+8CN - +O2+2H2O→4Ag(CN) 2- +4OH- (5)

[0229] It may also be possible to solubilize platinum group metals (PGMs) using sodium cyanide. Indeed, under pressure, soluble complexes are formed according to the following equation: 2Pt(s)+8NaCN+O2+2H2O→2Na2[Pt(CN)4]+4NaOH (6) 2Pd(s)+8NaCN+O2+2H2O→2Na2[Pd(CN)4]+4NaOH (7)

[0230] Gold can also be dissolved by complexation with thiosulfate solution in the presence of copper and ammonium ions according to the following equation: Au(s)+5S2O3 2- +Cu(NH3)4 2+ →Au(S2O3)2 3- +4NH3+Cu(S2O3)3 5- (8)

[0231] In some implementations, leaching of precious metals may be carried out using a third leach solution including an inorganic acid such as sulfuric acid, a thiourea solution (leaching agent concentration of 0.10-0.50 parts leaching agent per gram of metal-depleted residual solids), and a ferric ion salt (oxidizing agent 0.00-0.10 parts oxidizing agent per gram of metal-depleted residual solids). The third leach solution may have an acid concentration of 0.1-0.5N. Leaching of precious metals may be carried out according to a solids content of 50-200 grams of metal-depleted residual solids per liter of third leach solution.

[0232] For example, an acid concentration of 0.2N, a thiourea concentration of 0.25 g / g, and 0.06 g Fe per g of metal-reduced residual solids 3+ may be used at a solids content of approximately 130 g / L and a reaction time of 120 minutes. The temperature may be maintained between 5 and 95°C during the leaching of the precious metals, with a recommended temperature being 20°C.

[0233] Referring to Figure 3, the process further includes a third solid-liquid separation (102) for separating the third solid-liquid mixture (96) into a precious metal enriched leach solution (98) and a metal concentrate (100).

[0234] Selective extraction of REEs, base metals, and precious metals (if present) MnO2 recovery The metal-depleted residual solids resulting from the co-leaching step of REEs and base metals from the metal powder may consist primarily of manganese oxide (MnO2) concentrate (85% or greater purity), which can be reused industrially. This final solids may therefore be dried before being recycled industrially or sold as a concentrated manganese source.

[0235] REE recovery The REE-enriched leachate may be further processed to extract and isolate the extracted REE-containing components.

[0236] In one embodiment, extracting the REEs may include first precipitating the REE hydroxides and / or REE sulfates by adding a base solution to the REE-enriched leachate to produce a fourth solid-liquid mixture comprising the REE-bearing component and a secondary leachate, the REE-bearing component comprising the REE hydroxides and / or REE sulfates. The REE-bearing component and the secondary leachate may then be separated by solid-liquid separation.

[0237] In another embodiment, referring to FIG. 4 , extracting the REEs may include precipitating (106) REE hydroxides and / or REE sulfates by adding a base solution (108) to the metal-enriched leachate (104) to form a fourth solid-liquid mixture (110), separating (112) the REE hydroxides and / or REE sulfates (114) from the fourth solid-liquid mixture (110) via solid-liquid separation (112), redissolving the REE hydroxides and / or REE sulfates (114) in an acid solution (120) to form an REE-containing solution (122), and precipitating (124) the REE oxalates by adding oxalic acid and / or oxalate salts (126) to the REE-containing solution (122) to form a fifth solid-liquid mixture (128) comprising the REE oxalates. The REE oxalate may then be separated (130) from the fifth solid-liquid mixture (128) as an REE-containing component (132).

[0238] In yet another embodiment, extracting the REEs comprises precipitating the REE oxalates by adding sulfate to the REE-enriched leachate to produce a second solid-liquid mixture comprising the REE oxalates and a second leachate, and separating the REE oxalates and the second leachate as the REE-containing components from the fourth solid-liquid mixture via solid-liquid separation.

[0239] Precipitation of REE hydroxides and / or REE sulfates can occur when the pH is increased to a value higher than 1.2. The pH can be increased to 1.5 by adding a base solution such as sodium hydroxide. Potassium hydroxide (KOH), calcium hydroxide (Ca(OH)2), and magnesium hydroxide (Mg(OH)2) can also be used as base solutions. The fourth solid-liquid mixture can be mixed for 5 to 60 minutes, preferably 10 minutes. The temperature of the fourth solid-liquid mixture can be maintained at 5 to 95°C during the precipitation process, with a recommended temperature being 20°C.

[0240] Precipitation of REE hydroxides and / or REE sulfates may be preferred when using sulfuric acid as the acid solution. Precipitation of REE oxalates may also be facilitated by the addition of sulfates such as sodium sulfate or potassium sulfate or any mixture thereof.

[0241] The fifth solid-liquid mixture (or the second solid-liquid mixture) may be mixed for 5 to 60 minutes, preferably 10 minutes. The temperature of the fifth solid-liquid mixture (or the second solid-liquid mixture) may be maintained between 5 and 95°C during precipitation of the REE oxalates, with a recommended temperature being 20°C. The pH of the fifth solid-liquid mixture (or the second solid-liquid mixture) is maintained between 0.70 and 0.80, preferably at a pH of about 0.75. The REE oxalates may be recovered by solid-liquid separation. The purity of the resulting REE oxide concentrate may be 90% or higher.

[0242] 4, the REE-containing component (132) may also be converted to REE oxides (concentrate) (138) by calcination (136). Optionally, calcination (136) may be carried out at a calcination temperature of 700°C to 800°C for a duration of 2 to 4 hours.

[0243] The REE sulfates (114), REE oxalates (132), and / or REE oxides (138) produced during these processes can be used commercially or sold as a concentrated source of rare earth elements.

[0244] It should be noted that the process of the present invention allows for the combination of streams produced by different sequences of steps. For example, a metal and REE concentrated leachate (52) produced by co-leaching of base metals and REEs as shown in Figure 2 can be combined with an REE concentrated leachate (74) produced by selective leaching of REEs as shown in Figure 3 to form a leachate (104) as shown in Figure 4, thereby optimizing the efficiency of the overall process.

[0245] Basic metal recovery The process further includes selectively extracting base metals from at least one of the base-metal-enriched leachate and the secondary leachate to produce a plurality of extracted components, each extracting component comprising at least one of the base metals. Note that extracting REEs from the REE-enriched leachate produces a secondary leachate. When an REE-enriched leachate is produced by simultaneous leaching of REEs and base metals, such secondary leachate may contain base metals. Alternatively, when leaching of REEs and base metals is selective, i.e., when the e-waste comprises electronic waste, the base metals may rather be contained in the base-metal-enriched leachate.

[0246] In some embodiments, as seen in Figures 5 and 6, the process may include combining the base metal-enriched leachate (86) with the secondary leachate (116) to form a metal-enriched leachate (140). Selective extraction of base metals may then be performed from the metal-enriched leachate (140).

[0247] Referring to FIGS. 5 and 12, the basic metals may include copper, zinc, cadmium, manganese, cobalt, nickel, and lithium, and the multiple extracted components sequentially produced by the selective extraction of copper, zinc, cadmium, manganese, cobalt, nickel, and lithium include metallic copper (Cu), metallic zinc (Zn), metallic cadmium (Cd), manganese carbonate concentrate (MnCO), cobalt oxide concentrate (CoO), nickel oxide concentrate (NiO), lithium carbonate (LiCO), and lithium phosphate (LiPO).

[0248] It should be noted that selectively extracting the base metals can include sequentially extracting each one of the base metals to produce each extract component containing one base metal. Further details regarding exemplary extraction of each base metal are provided.

[0249] Copper recovery As used herein, the abbreviation "S / L ratio" means the solids / liquid ratio. As used herein, the abbreviation "O / A ratio" means the organic to aqueous ratio.

[0250] Referring to Figures 5 and 6, the first solvent extraction step (142) can consist of copper separation. The leachate (140) is first adjusted to a pH of about 2.0 by adding a base (144), preferably sodium hydroxide. The leachate (140) is also contacted with an organic solvent (146), which can be, for example, approximately 45% (v / v) of a mixed aldoxime-ketoxime reagent (e.g., LIX 84-I) and 55% (v / v) kerosene. A two-stage organic solvent extraction (142) with a 1:1 (v / v) organic phase (154) / aqueous phase (152) (O / A) ratio is typically required to completely extract copper from the aqueous phase (152). The temperature of the extraction step is maintained at about 20-60°C, preferably about 20°C. The organic and aqueous phases are separated by S / L separation (156). Copper is then selectively stripped (158) from the organic phase (154) by adding a solution of H2SO4 (2N) (160) in an O / A ratio of 1:1 (v / v). Most of the copper present in the organic phase (154) is stripped in the first step. The stripped solution (162) obtained from the stripping step (158) is recycled to the next cycle. The retention time for extraction (142) and stripping (158) is fixed at approximately 10 minutes for all steps. The copper present in the first stripping solution (164) is then electrodeposited (166) at a pH of about 2.0. During electrodeposition (166), the pH of the solution is maintained constant by the addition of a base (168), such as sodium hydroxide. The temperature in the electrodeposition step (166) is maintained at about 20-40°C, preferably about 20°C. Copper metal (170) having a purity of at least 98% can be recovered. An effluent (172) is produced as a by-product of the electrodeposition, which can be recycled.

[0251] Stainless steel or copper can be used as the cathode material, while Ti / IrO2 can be used as the anode material. 2 A current density fixed at 0.25 makes it possible to obtain a high removal yield of copper that deposits on the cathode in the form of metallic copper.

[0252] Zinc recovery Referring to Figure 7, the Cu-depleted solution (152) obtained after Cu removal can then be processed by solvent extraction (174) to concentrate the zinc. The leachate (152) is first adjusted to a pH of approximately 2.0-2.5 by adding a base (176), preferably sodium hydroxide. The leachate (170) is also contacted with an organic solvent (178) consisting of approximately 20-30% (v / v) bis(2,4,4-trimethylpentyl)phosphinic acid (Cyanex 272), 2% (v / v) tributyl phosphate (TBP), and 68-78% (v / v) kerosene. A two-stage organic solvent extraction (174) with a 2:1 (v / v) organic phase (182) / aqueous phase (184) (O / A) ratio is typically required to completely extract the zinc from the aqueous solution. The temperature of the extraction step (174) is maintained at about 40-60°C, preferably about 50°C. The organic phase (182) and aqueous phase (184) are separated by S / L separation (186). Zinc is selectively stripped (188) from the organic phase (182) by adding a solution (190) of H2SO4 (0.2N) at an O / A ratio of 2:1 (v / v). Most of the zinc present in the organic phase (182) is stripped in a single step (188). Residual iron (200) present in the second organic phase (192) can be stripped (194) by adding a more concentrated solution (196) of H2SO4 (e.g., 2.0N) at an O / A ratio of 2:1 (v / v). The stripped solution (198) obtained from the second stripping step (194) is recycled to the next cycle. The retention time for extraction and stripping is fixed at approximately 10 minutes for all steps.

[0253] The zinc present in the first stripping solution (202) is then electrodeposited (204) at a pH of about 2.0. During electrodeposition (204), the pH of the solution is maintained constant by the addition of a base (206), such as sodium hydroxide. The temperature of the electrodeposition step (204) is maintained at about 40-60°C, preferably about 50°C. Stainless steel or aluminum can be used as the cathode material, while Ti / IrO2 or Pb / Ag can be used as the anode material. The current is about 250-750 A / m with a hold time of approximately 2-3 hours. 2 The current density, which can be fixed at 0.5 V, results in a high removal yield of zinc, which deposits on the cathode in the form of metallic zinc (208). The purity of the metallic zinc (208) is typically at least 98%. A separate effluent (210) is produced as a by-product of the electrodeposition, which can be recycled.

[0254] Cadmium Recovery Referring to Figure 8, the next step following the extraction of zinc is the recovery of cadmium from the Zn-depleted aqueous solution (184). The pH of this solution (184) is first adjusted to about 2.7-2.9 by adding a base (214) before mixing with an organic solvent (216). The organic solvent (216) may contain, for example, approximately 30% (v / v) di-(2-ethylhexyl)phosphoric acid (DEHPA), 5% (v / v) TBP, and 65% (v / v) kerosene. A two-stage organic solvent extraction (212) with an O / A ratio of 2:1 (v / v) is usually required to completely extract cadmium from the aqueous solution. The temperature of the extraction step (212) is maintained at about 40-60°C, preferably about 50°C. The equilibrium pH is adjusted to 2.2-2.9 during the extraction. Cd and Mn are co-extracted and transferred to the organic phase (224) of the extraction solution (218). A separation (220) of the organic phase (224) and an aqueous phase (222) can be performed, with the aqueous phase (222) referred to as the Cu, Zn, Cd, Mn-depleted solution (222). After separation (220), the organic phase (224) can undergo a first scrubbing phase (226) used to remove major impurities, such as Ni and Co, from the organic phase (224) using an acid solution (225), thereby forming a purified organic phase (228). An additional stripped solution (230) can form part of the Cu, Zn, Cd, Mn-depleted solution (222). The O / A ratio of the scrubbing (226) can be equal to 20:1 (v / v), and its initial pH is fixed at about 2.3. The Cd and Mn present in the purified organic phase (228) can then be further stripped by the addition of a 0.5-0.6 N H2SO4 solution (234) with an O / A ratio of 4:1 (v / v). The reaction times for the extraction steps (212, 226, 232), including the scrubbing step (226) and the stripping step (232), are fixed at 10 minutes for all of these steps.

[0255] The stripped aqueous solution (236) can then be transferred to an electrolysis compartment, where Cd is selectively electrodeposited (240) from the aqueous solution (236) while other metals (Mn and trace amounts of Ni, Co, Zn) remain in the aqueous solution. The selective electrodeposition / electrowinning (240) of Cd occurs at a pH of about 2.0 and a current of about 360-370 A / m 2 The electrodeposition process was carried out at a fixed current density of 0.05 V and a holding time of approximately 2 to 4 hours. The temperature of the electrodeposition step (240) is maintained at approximately 40 to 60°C, preferably at approximately 50°C. During electrodeposition, the pH of the solution (236) is kept constant by the addition of a base (242) such as sodium hydroxide. Stainless steel can be used as the cathode material, while Ti / IrO2 can be used as the anode material. The purity of the metallic cadmium (244) deposited at the cathode can be at least 98%. A manganese-enriched solution (246) can also be recovered.

[0256] In other implementations, cadmium can be recovered by precipitation in the form of cadmium sulfide (CdS) instead of an electrodeposition process. This CdS precipitation occurs after raising the pH to approximately 4.0 by adding a base to the stripped aqueous solution 236 and by adding a sulfide salt such as NaS or NaHS. A solids / liquid separation process is then used to separate the CdS precipitate from the manganese-rich solution.

[0257] Manganese recovery After cadmium electrodeposition, the remaining MnSO solution (246) is transferred to a precipitation step (250), as shown in Figure 9. The solution is adjusted to a pH of approximately 7.0 by the addition of a base (252) (e.g., NaOH or KOH solution) followed by the addition of a carbonate (254) (e.g., NaCO or KCO). The manganese present in the MnSO solution (246) precipitates at a pH of approximately 8-9. The solution is mixed for 15-240 minutes. The resulting MnCO precipitate (256) is then recovered by solid-liquid separation (260), and a separate effluent (258) is also recovered separately for recycling. The solids (256) may be washed with water to remove impurities before dewatering and drying (not shown in Figure 9). The final product can be recycled industrially or sold as a concentrated manganese source. The purity of the resulting MnCO3(256) can be at least 94%.

[0258] Cobalt Recovery Referring to Figure 10, the next step in the process sequence is the recovery of cobalt from the Zn, Cd, and Mn-depleted solution (222). The pH of this solution (222) is first adjusted to about 5.2-5.5 by adding a base (262) before mixing with an organic solvent (264). The organic solvent (264) contains approximately 10% (v / v) bis(2,4,4-trimethylpentyl)phosphinic acid (Cyanex 272), 2% (v / v) TBP, and 88% (v / v) kerosene. A single organic solvent extraction (266) with an O / A ratio of 1:2 (v / v) is typically required to completely extract Co and Ni from the aqueous phase (270). The temperature of the extraction step (266) is maintained at about 40-60°C, preferably about 50°C. Co and Ni are co-extracted and transferred to the organic phase (272). Separation (268) of an organic phase (272) and an aqueous phase (270) is further performed, with the aqueous phase (270) being referred to as a Cu, Zn, Cd, Mn, Co-depleted solution (270). After separation (268), the organic phase (272) may undergo a first scrubbing (274) performed by adding a scrubbing agent (276), such as CoSO, to remove Ni from the organic phase (272) and form a depleted organic phase (278). The secondary aqueous phase (280) recovered from scrubbing (274) may be used as part of the Cu, Zn, Cd, Mn, Co-depleted solution (270). The O / A ratio of scrubbing (274) was equal to 4:1 (v / v). The Co present in the organic phase (278) can then be stripped (282) by the addition of a 0.25 N HSO solution (284) with an O / A ratio of 2:1 (v / v). The reaction time for the extraction steps (266, 274, 282), including the scrubbing step (274) and the stripping step (282), is fixed at 10 minutes for all of these steps.

[0259] The Co present in the resulting CoSO solution (286) is then removed by cobalt oxalate precipitation (290). This is accomplished by the addition of a precipitating agent (292), which may be oxalate or oxalic acid. The resulting cobalt oxalate precipitate (296) and additional effluent for recycle (298) are then recovered by solid-liquid separation (294). The cobalt oxalate precipitate (296) may be further converted to cobalt oxide (300) by calcination (298). The CoO concentrate (300) produced according to these steps may be used industrially or sold as a concentrated cobalt source. The purity of the resulting CoO may be at least 90%.

[0260] Nickel recovery Referring to FIG. 11, the Ni present in the resulting Zn, Cd, Mn, and Co-depleted solution (270) is then removed by precipitation of Ni oxalate (302). This is accomplished by the addition of a precipitating agent (304), which may be an oxalate salt or oxalic acid. The pH may also be adjusted via the addition of a base (306). The resulting solution is mixed for 15 to 240 minutes. The Ni oxalate precipitate (310) and tertiary leachate (312) are then recovered by solid-liquid separation (308). The Ni oxalate precipitate (310) is then converted to Ni oxide (NiO concentrate) (316) by calcination (314). The recovered NiO concentrate (316) can be used industrially or sold as a concentrated nickel source. The purity of the resulting NiO can be at least 90%.

[0261] Lithium recovery Referring to Figure 12, at the end of the solvent extraction series, lithium present in the aqueous fraction (tertiary leachate (312)) is recovered by selective precipitation in the form of lithium carbonate (Li2CO3) and lithium phosphate (Li3PO4). Before lithium separation is performed, it is preferable to remove impurities by a precipitation step (318) of the metals still present by the formation of metal hydroxides at a pH of about 7-10. The pH is adjusted by the addition of a base solution (320), such as NaOH, KOH, or Ca(OH)2, to the tertiary leachate (312). The precipitate (324) (metal hydroxide sludge) is recovered by solid-liquid separation (322), thereby allowing further recovery of a lithium-containing solution (326) (e.g., a lithium sulfate solution).

[0262] The resulting lithium-containing solution (326) is then treated by adding a carbonate or a concentrated solution of a carbonate (328), preferably sodium carbonate (NaCO), to induce precipitation of lithium carbonate (330). The amount of carbonate (328) added must be sufficient to precipitate all of the lithium according to the following reaction: Li2SO4(aq)+Na2CO3(aq)→Li2CO3(s)+Na2SO4(aq) (9)

[0263] The temperature of the solution (326) is adjusted between 0 and 100°C, preferably at about 50°C. Note that the solubility of lithium carbonate decreases with increasing temperature, i.e., 15.4 g / L (0°C), 12.8 g / L (25°C), 10.7 g / L (50°C), and 7.2 g / L (100°C). The solubility of sodium carbonate increases between 0 and 40°C and then remains nearly constant at approximately 300 g / L up to 100°C. The pH of the solution (326) should preferably be maintained between approximately 9.0 and 10.0. The solution is mixed for 15 to 240 minutes. Sufficient mixing time must be maintained to obtain the formation of Li2CO3 crystals. The formed lithium carbonate (334) is then recovered by solid-liquid separation (332). This solid (334) can be washed with water to remove some of the impurities, then dried and either recycled industrially or sold as a concentrated lithium source.

[0264] After the lithium carbonate precipitation step (330), the lithium remaining in solution (336) (referred to as lithium-depleted solution (336)) can be recovered by precipitation of lithium phosphate (340) after addition of a phosphate salt (342), preferably sodium phosphate (NaPO). The desired reaction is as follows: 3Li2SO4(aq)+2Na3PO4(s)→2Li3PO4(s)+3Na2SO4(aq) (10)

[0265] The solubility of Li3PO4 is lower than that of Li2CO3, with a value of approximately 390 mg / L at alkaline pH.

[0266] 12, the LiPO precipitate (346) thus formed is recovered by S / L separation techniques (344). This solid (346) can also be washed and dried and then recycled industrially or sold as a concentrated lithium source. If the lithium concentration in the solution at the end of the solvent extraction step is too low, it may be possible to proceed directly to the lithium phosphate precipitation step without going through the lithium carbonate precipitation step.

[0267] Precious metal recovery The process may further include extracting precious metals from the precious metal concentrated leachate to produce a precious metal concentrate and a recyclable effluent.

[0268] In some embodiments, the extraction of precious metals from the precious metal enriched leachate comprises using at least one of activated carbon adsorption, cementation, ion exchange, or electrodeposition.

[0269] For example, referring to Figure 13, adsorption (350) of the precious metals contained in the precious metal enriched leachate (98) can be carried out on an activated carbon support (352) (e.g., activated coal). After further solid-liquid separation (354), the precious metal concentrate (356) and the final effluent (358) for recycling (358) can be separately recovered. Other conventional recovery techniques used in the mining industry can also be used to recover the precious metal concentrate.

[0270] It should be noted that the final solids (metal concentrate) (100) obtained at the end of precious metal leaching, as shown in Figure 3, consists mainly of a palladium concentrate (75% or higher purity) that can be reused industrially. Therefore, this residue may be dried before being recycled industrially or sold as a concentrated palladium source. [Example]

[0271] Example 1 - Leaching of metals and REEs from a mixture of used batteries The collected spent batteries were frozen using liquid nitrogen and then crushed to remove steel castings. The fine particles were screened through a 1-2 mm opening sieve, dried at 60°C, and then crushed. The fine particles were mixed for 45 minutes at ambient temperature (20°C) and 80°C with different sulfuric acid concentrations and solids concentrations of 100 g of powder per liter of acid solution. After processing, the remaining powder was separated from the solution by vacuum filtration. The percentages of solubilized metals and solubilized REEs presented in Table 3 (20°C) and Table 4 (80°C) were established based on their concentrations measured in solution.

[0272] The results show increased solubilization of metals and REEs with increasing sulfuric acid content. High solubilization yields are obtained using 2.0N or 4.0N H2SO4. The solubilization yields obtained at high temperatures are approximately 10-20% higher than those at room temperature for the metals and REEs of interest. Thus, for battery recycling, greater than 70% solubilization of all metals and REEs of economic interest (Cd, Co, Ce, La, Li, Mn, Nd, Ni, Y, and Zn) was obtained after a single leaching step at 80 °C in the presence of H2SO4 (2N), except for manganese (40%), which exists primarily in the form of manganese dioxide (MnO2) and is practically insoluble under moderately acidic conditions.

[0273] It should be noted that the solubilization yield can be enhanced by applying more than one leaching step. The results show that it is also possible to remove many of the potassium and sodium ions by washing the spent battery powder before the acid leaching step. This washing can be carried out at room temperature as well as at elevated temperatures (e.g., 80°C).

[0274] [Table 3] Leaching time = 45 min, solid concentration = 100 g / L, one leaching step.

[0275] [Table 4] Leaching time = 45 min, solid concentration = 100 g / L, one leaching step.

[0276] Example 2 - Leaching of REEs and metals from smartphone waste Electronics powder manufacturing The collected samples are disassembled into their different components: plastic, printed circuit board assembly, battery, and screen. The printed circuit board assembly is then shredded, crushed, and sieved. After sieving, a powder with a particle size of less than 425 μm is obtained. After being divided into four parts, the resulting electronics powder is leached in an acid solution.

[0277] Solubilization of REEs and metals Solubilization of REEs The sample powder is leached with a solution of a strong inorganic acid (hydrochloric acid, nitric acid, sulfuric acid), preferably sulfuric acid (0.5-5N). Preferably, a 0.4N sulfuric acid concentration is used with a solids content of approximately 100 g / L and a reaction time of 30 minutes. The temperature of the mixture is maintained between 20 and 80°C during the leaching process, with a recommended temperature being 20°C. After the leaching process, the residual solids and the acidic solution containing the REEs are separated by a solid-liquid separation process, such as filtration, centrifugation, or sedimentation.

[0278] The fine particles were mixed for 45 minutes at ambient temperature (20°C) and 80°C with different sulfuric acid concentrations and solids concentrations of 100 g of powder per liter of acid solution. After processing, the residual powder was separated from the solution by vacuum filtration. The percentages of solubilized REEs presented in Table 5 were established based on their concentrations measured in solution.

[0279] [Table 5] Leaching time = 30 min, solid concentration = 100 g / L, one leaching step.

[0280] Copper solubilization The residue from rare earth element leaching is used for leaching of basic metals. For this purpose, a dilute inorganic acid solution (2.0N-4.0N) and a strong oxidizing agent such as hydrogen peroxide (0.33-1.33 g H2O2 per gram of powder) are used. The amount of oxidizing agent to be added can be determined based on stoichiometry by knowing the metal (mainly copper) content present in the powder mixture. A mixture of powder and acid solution with a solids content of 100-150 g per liter of solution is mixed for 30-180 minutes. Preferably, a sulfuric acid concentration of 2.0N and 0.67 g H2O2 per gram of powder are used, with a solids content of approximately 100 g / L and a reaction time of 180 minutes. The temperature of the mixture is maintained between 20-80°C during the leaching process, with 80°C being the recommended temperature. After the leaching process, the residual solids and the metal-containing acid solution are separated using a standard S / L separation process. According to our experiments, 1 L of leaching solution consisted of Cu (25.2 g) and Ni (1.4 g).

[0281] Solubilization of precious metals Solubilization of precious metals can be achieved by forming soluble complexes with thiourea, thiosulfate, or cyanide solutions. Thus, precious metal leaching is performed using sulfuric acid (0.2 N), thiourea solution (0.15-0.25 g / g), and ferric ion salts (0.05-0.06 g Fe per 1 g of powder). 3+ A mixture of powder and thiourea solution having a solids content of 60-131 g per liter of solution is mixed for 30-120 minutes. Preferably, a thiourea concentration of 0.25 g / g and 0.06 g Fe per gram of powder are used. 3+ is used with a solids content of approximately 131 g / L and a reaction time of 120 minutes. The temperature of the mixture is maintained between 18 and 25°C during the leaching process, with a recommended temperature of 20°C. After the leaching process, the residual solids and the acidic solution containing the precious metals are separated by a standard S / L separation process. In our experiments, 1 L of leaching solution consisted of Ag (5.7 g) and Ag (0.9 g).

Claims

1. 1. A process for treating electronic waste including at least one of used batteries and electronic waste, the process comprising: a) fragmenting the e-waste to reduce its size and generate fragmented waste comprising metal powder; b) recovering the metal powder from the fragmented waste, the metal powder comprising rare earth elements (REE) and basic metals; c) leaching the REE and the base metal from the metal powder to produce a residual metal-depleted solids and at least one of a REE-enriched leachate and a base metal-enriched leachate; d) extracting the REE from the REE-enriched leachate to produce a REE-containing component and a secondary leachate; e) selectively extracting the base metals from at least one of the base-metal-enriched leachate and the secondary leachate to produce a plurality of extract components, each extract component comprising at least one of the base metals.

2. 2. The process of claim 1, wherein the basic metal comprises Cd, Co, Cu, Li, Mn, Ni, Zn, or any combination thereof.

3. 3. The process of claim 1 or 2, wherein the REE comprises La, Nd, Sm, Ce, Pr, Y, Tb, Er, Sc, Gd, Eu, Dy, or any combination thereof.

4. 4. The process of claim 1, wherein recovering the metal powder comprises separating the fragmented waste into coarse and fine fractions, the fine fractions comprising the metal powder.

5. 5. The process of claim 4, wherein the fine fraction further comprises a low-density material comprising at least one of plastic and paper, and recovering the metal powder further comprises separating the low-density material and metal fragments from the fine fraction using a gravity-based method.

6. leaching the REE and the base metal from the metal powder contacting the metal powder with a leach solution to solubilize at least a portion of both the REE and the base metal to produce a first solid-liquid mixture comprising the REE-enriched leachate and metal-depleted residual solids; and separating the REE-enriched leachate and the metal-depleted residual solids by solid-liquid separation, thereby simultaneously leaching the REE and the base metals; The process of any one of claims 1 to 5, wherein the REE-enriched leachate further comprises the basic metal.

7. The process of claim 6 , wherein the leaching solution comprises an inorganic acid.

8. 8. The process of claim 7, wherein the leach solution has an acid concentration of 0.5N to 5N.

9. 9. The process of any one of claims 6 to 8, wherein the ratio of the weight of the metal powder to the volume of the leach solution is from 50 g / L to 200 g / L.

10. The process according to any one of claims 6 to 9, wherein the electronic waste consists of the used batteries.

11. 11. The process of any one of claims 6 to 10, wherein the leaching solution further comprises hydrogen peroxide, sodium metabisulfite, or a combination thereof.

12. The process of any one of claims 1 to 5, wherein the leaching of the REE and the base metal from the metal powder comprises selectively leaching the REE and the base metal.

13. 13. The process of claim 12, wherein selectively leaching the REE and the base metal comprises leaching the REE before leaching the base metal.

14. The leaching of the REE contacting the metal powder with a first leach solution to solubilize at least a portion of the REE and produce a first solid-liquid mixture comprising the REE-enriched leachate and a REE-depleted residual solids; and separating the REE-enriched leachate and the REE-depleted residual solids by solid-liquid separation.

15. 15. The process of claim 14, wherein the first leach solution comprises an inorganic acid.

16. 16. The process of claim 14 or 15, wherein the first leach solution has an acid concentration of from 0.2N to 1N.

17. 17. The process of any one of claims 14 to 16, wherein the ratio of the weight of the metal powder to the volume of the first leach solution is from 50 g / L to 200 g / L.

18. The leaching of the basic metal comprises: contacting the REE-depleted residual solids with a second leach solution to solubilize at least a portion of the base metals and produce a second solid-liquid mixture comprising the base metal-enriched leach solution and the metal-depleted residual solids; and separating the basic metal-enriched leachate and the metal-depleted residual solids by solid-liquid separation.

19. 20. The process of claim 18, wherein the second leach solution comprises an inorganic acid and an oxidizing agent.

20. 20. The process of claim 19, wherein the oxidizing agent comprises hydrogen peroxide.

21. 21. The process of claim 19 or 20, wherein the second leach solution comprises an oxidizing agent in an amount that is stoichiometrically sufficient to oxidize the base metal present in metallic form in the metal powder.

22. 22. The process of any one of claims 19 to 21, wherein the second leach solution has an acid concentration of from 1N to 5N.

23. 23. The process of any one of claims 19 to 22, wherein the second leach solution has a weight ratio of the oxidizing agent to the REE-reduced residual solids of from 0.33 to 1.

33.

24. 24. The process of any one of claims 18 to 23, wherein the ratio of the weight of the REE-depleted residual solids to the volume of the second leach solution is from 50 g / L to 200 g / L.

25. 25. The process of any one of claims 1 to 24, wherein the metal powder further comprises a precious metal, and the process further comprises leaching the precious metal from the metal-depleted residual solids to produce a precious metal-enriched leachate and a metal concentrate.

26. said leaching of said precious metals comprising: contacting the metal-depleted residual solids with a third leach solution to solubilize at least a portion of the precious metals and produce a third solid-liquid mixture comprising the precious metal-enriched leach solution and the metal concentrate; and separating the precious metal enriched leach solution and the metal concentrate by solid-liquid separation.

27. 27. The process of claim 26, wherein the third leach solution comprises a thiourea solution, a thiosulfate solution, a cyanide solution, or any mixture thereof.

28. 28. The process of claim 26 or 27, wherein the third leach solution has a leachant concentration of 0.10 to 0.50 grams of leachant per gram of the metal-depleted residual solids.

29. 29. A process according to any one of claims 26 to 28, wherein the leaching of the precious metals is carried out according to a solids content of the metal-depleted residual solids of 50 to 200 grams per litre of the third leach solution.

30. 30. The process of any one of claims 25 to 29, further comprising extracting the precious metals from the precious metal concentrated leachate to produce a precious metal concentrate and a recyclable effluent.

31. 31. The process of claim 30, wherein the extraction of the precious metal from the precious metal enriched leach solution comprises using at least one of activated carbon adsorption, cementation, ion exchange, or electrodeposition.

32. 32. The process of claim 30 or 31, wherein the precious metal comprises gold (Au), silver (Ag), platinum group metals (PGM), and any combination thereof.

33. The process of any one of claims 14 to 32, wherein the electronic waste comprises the used batteries and the electronic waste.

34. The extraction of the REE comprises: adding a base solution to the REE-enriched leachate to precipitate REE hydroxides and / or REE sulfates to produce a fourth solid-liquid mixture comprising the REE-containing component and the secondary leachate, wherein the REE-containing component comprises the REE hydroxides and / or REE sulfates; and separating the REE-containing component and the secondary leachate by solid-liquid separation.

35. The extraction of the REE comprises: adding a base solution to the metal-enriched leach solution to precipitate REE hydroxides and / or REE sulfates to form a fourth solid-liquid mixture; separating the REE hydroxides and / or the REE sulfates from the fourth solid-liquid mixture via solid-liquid separation; redissolving the REE hydroxide and / or the REE sulfate in an acid solution to form a REE-containing solution; adding oxalic acid and / or an oxalate to the REE-containing solution to precipitate the REE-oxalate to form a fifth solid-liquid mixture comprising the REE-oxalate; and separating the REE oxalate as the REE-containing component from the fifth solid-liquid mixture.

36. The base solution may be sodium hydroxide (NaOH), potassium hydroxide (KOH), calcium hydroxide (Ca(OH) ２ ), and magnesium hydroxide (Mg(OH) ２ 36. The process of claim 34 or 35, wherein the compound is selected from the group consisting of:

37. 37. The process of any one of claims 34 to 36, wherein the base solution further comprises a sulfate salt.

38. The extraction of the REE comprises: adding sulfate to the REE-enriched leachate to precipitate REE oxalate to form a fourth solid-liquid mixture comprising the REE oxalate and the secondary leachate; and separating the REE oxalate and the secondary leachate as the REE-containing component from the fourth solid-liquid mixture via solid-liquid separation.

39. The sulfate is sodium sulfate (Na ２ SO ４ ), potassium sulfate (K ２ SO ４ ), or any mixture thereof.

40. 40. The process of any one of claims 34 to 39, further comprising calcining the REE-containing component to produce a REE oxide concentrate.

41. 41. The process of any one of claims 1 to 40, further comprising combining the basic metal-enriched leachate with the secondary leachate to form a metal-enriched leachate, and wherein selective extraction of the basic metal is performed from the metal-enriched leachate.

42. 42. The process of any one of claims 1 to 41, wherein selectively extracting the base metals comprises sequentially extracting each one of the base metals to produce each extract component comprising one base metal.

43. The basic metals include cadmium, cobalt, and lithium, and the extracted components produced by the selective extraction of the cadmium, cobalt, and lithium include metallic cadmium (Cd), cobalt oxide concentrate (CoO), and lithium carbonate (LiCO ３ ) and lithium phosphate (LiPO ４ 43. The process of any one of claims 1 to 42, comprising at least one of:

44. The basic metals include copper, zinc, cadmium, manganese, cobalt, nickel, and lithium, and the plurality of extracted components sequentially produced by the selective extraction of the copper, zinc, cadmium, manganese, cobalt, nickel, and lithium include metallic copper (Cu), metallic zinc (Zn), metallic cadmium (Cd), cadmium sulfide concentrate (CdS), manganese carbonate concentrate (MnCO ３ ), cobalt oxide concentrate (CoO), nickel oxide concentrate (NiO), lithium carbonate (LiCO ３ ), and lithium phosphate (Li ３ P.O. ４ 43. The process of any one of claims 1 to 42, comprising:

45. 45. The process of any one of claims 1 to 44, wherein the electronic waste comprises used batteries, and the process further comprises deactivating the used batteries before fragmenting the electronic waste.

46. 46. ​​The process of claim 45, wherein the inactivation of the used batteries comprises placing the used batteries in a saline or acid solution, freezing the used batteries using liquid nitrogen, or a combination thereof.

47. 47. The process of any one of claims 1 to 46, wherein the leaching of the REE and the base metal from the metal powder comprises carrying out multiple sequential acid leaching steps of the REE and the base metal.

48. 47. The process of any one of claims 1 to 46, wherein the leaching of the REE and the base metal from the metal powder comprises carrying out at least one acid leaching of the REE and base metal and at least one washing of the metal-depleted residual solids with water.

49. 1. A process for treating spent batteries containing basic metals and rare earth elements (REEs), the process comprising: a) deactivating the used battery; b) fragmenting the used batteries to reduce their size and generating fragmented waste comprising metal powder; c) recovering the metal powder from the fragmented waste, the metal powder comprising the REE and the basic metal; d) simultaneously leaching the REE and the base metal by contacting the metal powder with a leach solution to solubilize at least a portion of both the REE and the base metal and produce a first solid-liquid mixture comprising a leach solution and a metal-depleted residual solids, wherein the leach solution comprises both the REE and the base metal; e) separating the leachate and the metal-depleted residual solids by solid-liquid separation; f) extracting the REE from the leachate to produce a REE-containing component and a secondary leachate, the secondary leachate comprising the basic metal; g) selectively extracting the base metals from the secondary leach solution to produce a plurality of extract components, each extract component comprising at least one of the base metals.

50. 50. The process of claim 49, wherein the inactivation of the used battery comprises one of placing the used battery in a saline or acid solution, or freezing the used battery using liquid nitrogen.

51. 51. The process of claim 49 or 50, wherein the fragmentation of the used battery is carried out in a non-oxidizing environment.

52. 52. The process of claim 51 , wherein the non-oxidizing environment comprises a nitrogen atmosphere.

53. 53. The process of any one of claims 49 to 52, further comprising at least one feature of any one of claims 1 to 11 and 25 to 48.

54. The used battery is an alkaline Zn / MnO ２ 54. The process of any one of claims 1 to 53, wherein the battery is selected from the group consisting of a lithium ion battery, a zinc-carbon battery, a nickel-cadmium battery, a nickel-metal hydride battery, a nickel- ...

55. 55. The process of any one of claims 1 to 54, wherein the electronic waste comprises pieces of equipment selected from the group consisting of laptops, televisions, CRT monitors, LCD monitors, LED monitors, mobile phones, smart phones, electronic tablets, photovoltaic (PV) panels, conventional hard drives (HDDs), solid state drives (SSDs), and any combination thereof.

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