Method for recovering valuable metals

Abstract

The present invention provides a method for effectively recovering recyclable valuable metals and lithium from lithium-containing waste.

Description

Method for recovering valuable metals

[0001] The present invention relates to a method for recovering valuable metals.

[0002]

[0003] With the advancement of the IT industry and the increasing use of lithium-ion batteries, the volume of waste lithium-ion batteries and scrap generated during the manufacturing process is also increasing. The cathode, which accounts for more than 60% of the cost of lithium-ion batteries, is increasingly made of ternary oxides (nickel, cobalt, manganese), which offer superior battery performance compared to pure materials. These expensive metals are entirely imported, and the technology to recover these discarded metals is a national project.

[0004] The explosive demand for lithium-ion batteries and the shortening life cycles of small digital appliances have led to a sharp increase in lithium-ion battery emissions, leading to growing interest both domestically and internationally in lithium-ion battery processing and recycling. Reusing the waste materials of these batteries when they reach the end of their lifespan not only reduces environmental pollution but also reduces raw material costs during lithium-ion battery production. Accordingly, various research projects are ongoing to explore methods for recovering valuable metals and lithium from lithium-containing waste.

[0005]

[0006] The present invention provides a method for effectively recovering recyclable valuable metals and lithium from lithium-containing waste.

[0007] The purposes of the present invention are not limited to those mentioned above, and other unmentioned purposes and advantages of the present invention can be understood through the following description and will be more clearly understood through embodiments of the present invention. Furthermore, it will be readily apparent that the purposes and advantages of the present invention can be realized by the means and combinations thereof described in the specification.

[0008]

[0009] One embodiment of the present invention provides a method for recovering a valuable metal, comprising the steps of: preparing a first mixture including a powder obtained by pulverizing lithium-containing waste and a first flux; melting the first mixture in an arc furnace under reducing conditions to produce a first molten substance; separating a first slag and an alloy from the first molten substance; removing impurities from the alloy and producing a granulated alloy; preparing an electrolytic tank including an electrolyte, an anode and a cathode immersed in the electrolyte, and a semi-permeable membrane provided between the anode and the cathode; accommodating the granulated alloy in a receiving portion of the anode and leaching it through electrolysis; and filtering the leached electrolyte to obtain a solution containing the valuable metal.

[0010] According to one embodiment of the present invention, the lithium-containing waste may include at least one selected from the group consisting of waste lithium-ion batteries, waste electrodes, and waste positive electrode active materials.

[0011] According to one embodiment of the present invention, the first flux is Al2O 3, CaO, SiO2, and It may include at least one selected from the group consisting of Fe2O3.

[0012] According to one embodiment of the present invention, the arc furnace may be any one selected from the group consisting of a DC Electric Arc Furnace (DC-EAF), an AC Electric Arc Furnace (AC-EAF), a plasma arc furnace, and a combination thereof.

[0013] According to one embodiment of the present invention, the step of producing the first molten material can be performed by a direct current electric arc furnace under an output condition of 50 kW or more and 3,000 kW or less.

[0014] According to one embodiment of the present invention, the size of the granulated alloy may be 1 mm or more.

[0015] According to one embodiment of the present invention, the electrolysis can be performed under voltage conditions of 3.5 V or more and 15 V or less.

[0016] According to one embodiment of the present invention, the electrolysis can be performed under temperature conditions of 15°C or higher and 80°C or lower.

[0017] According to one embodiment of the present invention, the electrolyte solution includes an electrolyte, and the concentration of the electrolyte may be 10% or more and 30% or less.

[0018] According to one embodiment of the present invention, the step of obtaining a solution containing the precious metal may include heating the extracted electrolyte to 30°C or higher and 40°C or lower and filtering the obtained solution.

[0019] According to one embodiment of the present invention, the method for recovering valuable metal may further include a step of melting a second mixture including the first slag and the third flux to obtain third slag and fume; and a step of separating a lithium solution from the fume.

[0020] The solutions to the above problems are not exhaustive and may be combined with some embodiments of the present disclosure. The various features of the present invention and their corresponding advantages and effects can be understood in more detail by referring to the detailed description below.

[0021]

[0022] A method for recovering valuable metals according to one embodiment of the present invention has the advantage of being able to effectively and easily recover valuable metals from discarded lithium-containing waste.

[0023] A method for recovering valuable metals according to one embodiment of the present invention can increase the recovery rate of lithium while minimizing environmental pollution by using a small amount of land and environmentally friendly energy.

[0024] In addition to the aforementioned effects, the specific effects of the present invention are described below along with the specific details for implementing the invention. Furthermore, the effects of the present invention are not limited to the effects described above and can be readily achieved by the means and combinations thereof described in the specification.

[0025]

[0026] Figure 1 is a flow chart showing a method for recovering valuable metals according to one embodiment of the present invention.

[0027] Figure 2 illustrates a direct current electric arc furnace used in a method for recovering valuable metals according to one embodiment of the present invention.

[0028] Figure 3 is a schematic diagram of a wet scrubber used in a method for recovering valuable metals according to one embodiment of the present invention.

[0029] Figures 4a and 4b are drawings showing an electrolysis tank according to one embodiment of the present invention.

[0030] Figure 5 is a flowchart showing a method for recovering lithium in a method for recovering valuable metals according to one embodiment of the present invention.

[0031] FIG. 6 is a photograph of a second molten material produced in an embodiment of the present invention and a second slag removed from the second molten material.

[0032]

[0033] In this specification, singular expressions include plural expressions unless the context clearly indicates otherwise.

[0034] In this specification, expressions such as "first," "second," and "(S1)", "(S2)" can describe various components, regardless of order and / or importance, and do not limit the components. These expressions can be used to distinguish one component from another. For example, without departing from the scope of the present disclosure, the first component can be referred to as the second component, and similarly, the second component can also be referred to as the first component.

[0035] The numerical range indicated by the term "to" in this specification refers to a numerical range that includes the values ​​described before and after the term as the lower limit and the upper limit, respectively. When multiple numerical values ​​are disclosed as the upper and lower limits of an arbitrary numerical range, the numerical range disclosed in this specification can be understood as any numerical range that has any one of the multiple lower limit values ​​and any one of the multiple upper limit values ​​as the lower limit and the upper limit, respectively. For example, when a to b, or c to d is described in the specification, it can be understood that a or more and b or less, a or more and d or less, c or more and d or less, or c or more and b or less is described.

[0036] In the past, in the process of recovering lithium using lithium-containing waste, there was a problem in that the lithium recovery rate was low due to various process factors during the process when the pretreated lithium-containing waste was melted to recover lithium.

[0037]

[0038] One embodiment of the present invention provides a method for recovering a valuable metal, comprising the steps of: preparing a first mixture including a powder obtained by pulverizing lithium-containing waste and a first flux; melting the first mixture in an arc furnace under reducing conditions to produce a first molten substance; separating a first slag and an alloy from the first molten substance; removing impurities from the alloy and producing a granulated alloy; preparing an electrolytic cell including an electrolyte, an anode and a cathode immersed in the electrolyte, and a semi-permeable membrane provided between the anode and the cathode; receiving the granulated alloy in a receiving portion of the anode and electrolyzing it; and filtering the electrolyte to obtain a solution containing the valuable metal.

[0039] A method for recovering valuable metals according to one embodiment of the present invention has the advantage of effectively and easily recovering valuable metals from discarded lithium-containing waste. Furthermore, the method for recovering valuable metals can increase the lithium recovery rate while minimizing environmental pollution by utilizing less land and environmentally friendly energy. Furthermore, the method for recovering valuable metals can shorten the lithium recovery time and reuse alloys generated as by-products during the lithium recovery process.

[0040]

[0041] Figure 1 is a flowchart illustrating a method for recovering valuable metals according to one embodiment of the present invention. Hereinafter, with reference to Figure 1, the method for recovering valuable metals according to one embodiment of the present invention will be described in more detail.

[0042]

[0043] 1. Method for recovering valuable metals

[0044] (S1) A step of preparing a first mixture including a powder obtained by crushing lithium-containing waste and a first flux;

[0045] According to one embodiment of the present invention, the method for recovering valuable metals includes the step of preparing a first mixture including a powder obtained by crushing lithium-containing waste and a first flux to facilitate the extraction of lithium by crushing lithium-containing waste to increase the surface area (S1).

[0046] According to one embodiment of the present invention, the lithium-containing waste is not particularly limited and may include compounds containing lithium in various forms. For example, the lithium-containing waste may include at least one selected from the group consisting of waste lithium-ion batteries, waste electrodes, and waste cathode active materials. Here, the waste electrodes and waste cathode active materials may be separate components from the waste lithium-ion batteries, and may be electrodes or cathode active materials discarded separately from the waste lithium-ion batteries.

[0047] According to one embodiment of the present invention, the waste lithium-ion battery is a lithium-ion battery that has reached the end of its life or is discarded after use, and may specifically be a primary battery containing lithium, a secondary battery containing lithium, or an all-solid-state battery containing lithium. For example, the waste lithium-ion battery may be scrap or waste of the primary battery, secondary battery, or all-solid-state battery.

[0048] According to one embodiment of the present invention, the waste cathode active material is a cathode active material commonly used in the relevant technical field, and may specifically include a lithium transition metal oxide. For example, the lithium transition metal oxide may be Li x1 CoO2(0.5 <x1<1.3), Li x2 NiO2(0.5 <x2<1.3), Li x3 MnO2(0.5 <x3<1.3), Li x4 Mn2O4(0.5 <x4<1.3), Li x5 (Ni a1 Co b1 Mn c1)O2(0.5 <x5<1.3, 0<a1<1, 0<b1<1, 0<c1<1, a1+b1+c1=1), Li x6 Ni 1-y1 Co y1 O2(0.5 <x6<1.3, 0<y1<1), Li x7 Co 1-y2 Mn y2 O2(0.5 <x7<1.3, 0≤y2<1), Li x8 Ni 1-y3 Mn y3 O2(0.5 <x8<1.3, 0≤y3<1), Li x9 (Ni a2 Co b2 Mn c2 )O4(0.5 <x9<1.3, 0<a2<2, 0<b2<2, 0<c2<2, a2+b2+c2=2), Li x10 Mn 2-z1 Ni z1 O4(0.5 <x10<1.3, 0<z1<2), Li x11 Mn 2-z2 Co z2 O4(0.5 <x11<1.3, 0<z2<2), Li x12 CoPO4(0.5 <x12<1.3) 및 Li x13 FePO4(0.5 <x13<1.3)로 이루어진 군에서 선택되는 하나 이상을 포함할 수 있다.

[0049] In this specification, "powder" may be defined as a pulverized product obtained by pulverizing lithium-containing waste. The average particle diameter (D) of the powder 50) is not particularly limited, but may be specifically 3 mm or less, 2 mm or less, 1 mm or less, 0.1 mm or less, 0.01 mm or less, and specifically may be any one of the plurality of lower limits of 0.005 mm or more. In addition, the particle size of the powder may be adjusted to 100 mesh or less, 90 mesh or less, 80 mesh or less, 70 mesh or less, 60 mesh or less, 50 mesh or less, 40 mesh or less, 30 mesh or less, 20 mesh or less, or 10 mesh or less, and specifically may be any one of the plurality of lower limits of 1 mesh or more. For example, the average particle diameter may be calculated by dispersing target particles in a dispersion medium, introducing the particles into a commercially available laser diffraction particle size measuring device (e.g., Microtrac S3500), and measuring the difference in diffraction pattern according to particle size when the particles pass through a laser beam, thereby calculating the particle size distribution.

[0050] For example, if the lithium-containing waste is a waste lithium-ion battery or waste cathode active material, the powder may be black mass.

[0051] According to one embodiment of the present invention, the step (S1) may include a step of calcining the powder. Specifically, by calcining the powder, substances such as organic binders contained in the powder can be effectively removed, thereby significantly reducing the content of impurities other than those to be recovered. For example, the step of calcining the powder is not particularly limited, but may specifically include a step of calcining at a temperature of 600°C or higher and 900°C or lower for a time of 10 hours or higher and 14 hours or less.

[0052] According to one embodiment of the present invention, the first flux can lower the melting point of the first mixture and lower the viscosity of the first melt in the melting process described below to control fluidity, while at the same time easily inducing a reduction reaction of the precious metal contained in the powder.

[0053] According to one embodiment of the present invention, the content of the first flux may be 30 wt% or more and 50 wt% or less, 31 wt% or more and 49 wt% or less, 35 wt% or more and 48 wt% or less, 36 wt% or more and 47 wt% or less, 37 wt% or more and 46 wt% or less, 38 wt% or more and 45 wt% or less, 39 wt% or more and 44 wt% or less, 40 wt% or more and 43 wt% or less, 40 wt% or more and 42 wt% or less, or 40 wt% or more and 41 wt% or less, based on the total weight of the first mixture. By controlling the content of the first flux within the above numerical range, the viscosity of the first molten material can be lowered, effectively controlling the fluidity, and at the same time, easily inducing a reduction reaction of the precious metal contained in the powder.

[0054] According to one embodiment of the present invention, the first flux is not particularly limited, but is specifically Al2O 3, CaO, SiO2, and It may include at least one selected from the group consisting of Fe2O3, and specifically Al2O 3, It may contain CaO and SiO2.

[0055] According to one embodiment of the present invention, based on the total weight of the first mixture, the content of Al2O3 may be 1 wt% or more and 13 wt% or less, 2 wt% or more and 12 wt% or less, 3 wt% or more and 11 wt% or less, 4 wt% or more and 10 wt% or less, 5 wt% or more and 9 wt% or less, 6 wt% or more and 8 wt% or less, or 7 wt% or more and 8 wt% or less, the content of CaO may be 11 wt% or more and 23 wt% or less, 12 wt% or more and 22 wt% or less, 13 wt% or more and 21 wt% or less, 14 wt% or more and 20 wt% or less, 15 wt% or more and 19 wt% or less, 16 wt% or more and 18 wt% or less, or 17 wt% or more and 18 wt% or less, and the content of SiO2 may be 10 wt% or more and 22 wt% or less, 11 wt% or more and 21 wt% Below, it can be 12 wt% or more and 20 wt% or less, 13 wt% or more and 19 wt% or less, 14 wt% or more and 18 wt% or less, 15 wt% or more and 17 wt% or less, or 16 wt% or more and 17 wt% or less. By adjusting each composition in the first flux within the above numerical range, the viscosity of the first melt can be lowered, effectively controlling the fluidity, and at the same time, easily inducing a reduction reaction of the precious metal included in the powder.

[0056] According to one embodiment of the present invention, the first mixture may further include any one additive selected from the group consisting of transition metals, carbonates, and mixtures thereof. Specifically, the additive can further lower the melting point of the first mixture and effectively control the fluidity of the first melt in the melting process described below. For example, the transition metal is not particularly limited, but may specifically include Cu, and the carbonate is not particularly limited, but may specifically include any one selected from the group consisting of CaCO3, MgCO3, and mixtures thereof. For example, the content of the additive is not particularly limited, but specifically may be 3.0 wt% or less, 2.0 wt% or less, 1.0 wt% or less, or 0.5 wt% or less, and may be 0.1 wt% or more, based on the total weight of the first mixture.

[0057]

[0058] (S2) A step of reducing and melting the first mixture in an arc furnace under reducing conditions to produce a first melt;

[0059] In this specification, the "reduction condition" refers to a condition for forming a reducing atmosphere during the smelting process of the step (S2), and may be a condition for reducing a precious metal in the powder obtained by pulverizing the lithium-containing waste to produce an alloy to be described later. For example, the reduction condition is not particularly limited and may be formed by separately adding a reducing agent to the arc furnace or the first mixture, or may be implemented through an electrode (e.g., a carbon electrode) provided in the arc furnace without separately adding the reducing agent.

[0060] In this specification, “Arc furnace” is not particularly limited and may be a furnace that heats the first mixture using an electric spark generated when a high voltage difference is applied to positive and negative terminals.

[0061] In some embodiments of the present invention, the arc furnace may be any one selected from the group consisting of a DC Electric Arc Furnace (DC-EAF), an AC Electric Arc Furnace (AC-EAF), a plasma arc furnace, and combinations thereof, and specifically may be a DC Electric Arc Furnace (DC-EAF). According to some embodiments of the present invention, when a DC Electric Arc Furnace is used, a melting process can be effectively induced by using a higher current density and power than other types of arc furnaces, thereby increasing the recovery rate of lithium.

[0062] Meanwhile, a direct current electric arc furnace is described in detail with reference to Fig. 2.

[0063] Figure 2 illustrates a direct current electric arc furnace used in a method for recovering valuable metals according to one embodiment of the present invention.

[0064] Referring to FIG. 2, a direct current electric arc furnace (100) according to one embodiment of the present invention may include an electrode rod (10), a lower electrode (20), a separating section (30), an exhaust section (40), and a cooling section (50).

[0065] According to one embodiment of the present invention, the electrode rod (10) can generate an arc to heat and melt the first mixture. For example, the electrode rod (10) may be formed vertically (e.g., in the height direction) through an opening in the center of a roof of a direct current electric arc furnace (100). In one embodiment, the number of the electrode rods is not particularly limited, but may be specifically 1, 2 or more, or 3 or more, and 5 or less.

[0066] According to one embodiment of the present invention, the lower electrode (20) may be electrically connected to the electrode rod (10) to provide a path for current to flow. For example, the lower electrode (20) may be placed at the bottom of a direct current electric arc furnace (100), and specifically, may be placed at the bottom of an alloy, which is one of the melting results of the first mixture.

[0067] According to one embodiment of the present invention, the separation unit (30) can separate the first slag (S) and the alloy (A) generated in the DC electric arc furnace (100). Specifically, the separation unit (30) may include a first separation unit (30a) for separating the alloy (A) and a second separation unit (30b) for separating the first slag (S). For example, the separation unit (30) may be connected to both side walls of the DC electric arc furnace (100), and specifically, the first separation unit (30a) may be connected to one side wall of the DC electric arc furnace (100), and the second separation unit (30b) may be connected to the other side wall opposite to the one side wall.

[0068] According to one embodiment of the present invention, the exhaust section (40) can discharge gas generated in the DC electric arc furnace (100). For example, the exhaust section (40) can include a hole penetrating a portion of the roof of the DC electric arc furnace.

[0069] According to one embodiment of the present invention, the cooling unit (50) can inject cooling water or cooling air into the interior of the DC electric arc furnace (100) to prevent overheating of the DC electric arc furnace (100) and to induce safe operation at high temperatures. For example, the cooling unit (50) may include a first cooling unit (50a) that introduces cooling water into the interior of the DC electric arc furnace (100), and may include a second cooling unit (50b) that introduces cooling air into the interior of the DC electric arc furnace (100).

[0070] According to one embodiment of the present invention, the arc furnace includes a direct current electric arc furnace (100) equipped with an electrode rod (10), the electrode rod (10) includes a carbon electrode, and the reduction condition may be implemented by the carbon electrode. According to some embodiments of the present invention, since the electrode rod includes a carbon electrode, the valuable metal can be effectively reduced without adding a separate reducing agent, thereby easily obtaining an alloy to be described later. Accordingly, the cost can be reduced and the process efficiency can be improved, thereby improving the economic feasibility of the process, and at the same time, the recovery rate of lithium can be increased. For example, the carbon electrode may include graphite.

[0071] According to one embodiment of the present invention, the step (S2) may include a step performed by a direct current electric arc furnace under output conditions of 50 kW or more and 3,000 kW or less. Specifically, by adjusting the output conditions of the step (S2) within the above numerical range, the recovery rate of lithium can be further increased. Specifically, the output conditions of the step (S2) may be appropriately modified depending on the process scale.

[0072] According to one embodiment of the present invention, the current conditions of step (S2) may be appropriately modified depending on the process scale. For example, the current conditions may be 130 A or more and 150 A or less.

[0073]

[0074] (S3) A step of separating the first slag and alloy from the first molten material;

[0075] According to one embodiment of the present invention, the method for recovering the valuable metal includes a step (S3) of separating the first slag and the alloy from the first molten material to separate the first slag containing lithium and the alloy containing the valuable metal.

[0076] Specifically, the above step (S3) may be performed using a tapping process commonly used in the relevant technical field. For example, the separating member (30) or tapping hole commonly used in the relevant technical field may be used as a means to achieve the above step (S3).

[0077] Although FIG. 2 shows two separating parts (30), the technical idea of ​​the present invention is not limited thereto, and the number of separating parts (tapping holes) may be modified to two or more, three or more, four or more, or five or less.

[0078] For example, unlike the first slag, the alloy may not contain lithium.

[0079] For example, the composition of the alloy may vary depending on the composition of the lithium-containing waste. In one example, when the lithium-containing waste includes a LiNiCoMn cathode active material, the alloy may include all of Ni, Co, and Mn. In this case, the sum of the total contents of Ni, Co, and Mn based on the total weight of the alloy may be 85 wt% or more, 86 wt% or more, 87 wt% or more, 88 wt% or more, 89 wt% or more, or 90 wt% or more, and may be 95 wt% or less.

[0080]

[0081] 1-1. Method for removing impurities from alloys

[0082] Hereinafter, a method for recovering valuable metals according to one embodiment of the present invention will be described, in which impurities contained in an alloy are removed. Specifically, the method for recovering valuable metals may include, after step (S3), a process for removing impurities such as manganese and iron from the alloy.

[0083] (S4) A step of producing a second melt by reacting the above alloy with a second flux and oxygen;

[0084] According to one embodiment of the present invention, the method for recovering the valuable metal includes the step of (S4) reacting the alloy with a second flux and oxygen to produce a second molten material in order to remove impurities contained in the alloy.

[0085] According to one embodiment of the present invention, various furnaces commercially available in the relevant technical field can be used without particular limitation as a means for achieving the above step (S4). For example, a Top Blown Rotary Converter (TBRC), a Bottom Blown Converter, a Kaldo Converter, a Tilting Converter, etc., which can supply oxygen and remove impurities, can be used.

[0086] According to one embodiment of the present invention, in order to effectively remove impurities contained in the alloy, the alloy may be melted by supplying a second flux and oxygen together. Specifically, manganese (Mn) and iron (Fe) components contained in the alloy can be effectively removed. Through this, the contents of nickel (Ni) and cobalt (Co) components contained in a granulated alloy manufactured from the alloy can be effectively increased.

[0087] According to one embodiment of the present invention, the second flux is Al2O 3, CaO, SiO2, and It may include at least one selected from the group consisting of Fe2O3. Specifically, the second flux may include at least SiO2. In addition, the second flux may include Al2O 3, It may include CaO, and SiO2. By using the second flux including the above-mentioned components, manganese and iron, which are impurity components included in the alloy, can be effectively oxidized, and easily removed from the alloy in the form of slag.

[0088] According to one embodiment of the present invention, the amount of the second flux supplied to the alloy may be set according to the contents of manganese and iron contained in the alloy. Specifically, the amount of the second flux supplied may be set so that the ratio of the total moles of manganese and iron contained in the alloy to the moles of oxygen contained in the second flux is 1:1 to 1:5. When the amount of the second flux supplied is within the above-described range, manganese and iron, which are impurity components contained in the alloy, can be effectively oxidized. Through this, the contents of nickel and cobalt components contained in the alloy from which the second slag has been removed can be effectively increased.

[0089] According to one embodiment of the present invention, with respect to 100 parts by weight of the alloy, the amount of the second flux supplied may be 5 parts by weight or more and 15 parts by weight or less, 7 parts by weight or more and 13 parts by weight or less, 7 parts by weight or more and 9 parts by weight or less, 5 parts by weight or more and 10 parts by weight or less, or 8 parts by weight or more and 15 parts by weight or less. By adjusting the amount of the second flux supplied within the above-described range, manganese and iron, which are impurity components included in the alloy, can be effectively oxidized.

[0090] According to one embodiment of the present invention, the ratio of the components included in the second flux can be set in various ways. Specifically, the ratio of the components included in the second flux can be set in various ways within a range where the ratio of the total moles of manganese and iron included in the alloy and the moles of oxygen included in the second flux satisfies the above-described range. For example, the second flux may be Al2O 3, When containing CaO and SiO2, Al2O 3, The mixing ratio of CaO and SiO2 may be 0.1 to 10: 0.1 to 10: 0.1 to 10.

[0091] According to one embodiment of the present invention, the step of preparing the second melt may be performed at a temperature of 1,000°C or more and 2,000°C or less for a time of 2 minutes or more and 1 hour or less. Specifically, the temperature at which the step of preparing the second melt is performed may be 1,200°C or more and 1,800°C or less, 1,300°C or more and 1,600°C or less, 1,000°C or more and 1,500°C or less, or 1,400°C or more and 2,000°C or less. By controlling the temperature at which the step of preparing the second melt is performed within the above-described range, the second flux and oxygen can be effectively reacted with the alloy. Through this, the manganese and iron components included in the alloy are oxidized, so that manganese oxide and iron oxide can be easily removed from the alloy in the form of slag.

[0092] In addition, the time for which the step of producing the second melt is performed may be 2 minutes to 60 minutes, 2 minutes to 45 minutes, 2 minutes to 30 minutes, 5 minutes to 20 minutes, 5 minutes to 15 minutes, 2 minutes to 10 minutes, or 15 minutes to 30 minutes. When the time for which the step of producing the second melt is performed is within the above-mentioned range, the second flux and oxygen can be stably reacted with respect to the alloy, thereby effectively oxidizing manganese and iron included in the alloy.

[0093] According to one embodiment of the present invention, the step of producing the second molten material is performed at an oxygen partial pressure of 2 kgf / cm 2 More than 40 kgf / cm 2 It can be performed under the following conditions. Specifically, the partial pressure of oxygen supplied to the mixture containing the alloy and the second flux is 2 kgf / cm 2 More than 40 kgf / cm 2 Below 5 kgf / cm 2 More than 30 kgf / cm 2Below 5 kgf / cm 2 More than 20 kgf / cm 2 Below 7.5 kgf / cm 2 More than 18 kgf / cm 2 Below 9 kgf / cm 2 More than 15 kgf / cm 2 Below 5 kgf / cm 2 More than 15 kgf / cm 2 Below 8 kgf / cm 2 More than 12.5 kgf / cm 2 Below 20 kgf / cm 2 More than 40 kgf / cm 2 or less, or 25 kgf / cm 2 More than 35 kgf / cm 2 By controlling the partial pressure conditions of oxygen supplied to the alloy within the above-mentioned range, manganese and iron contained in the alloy can be effectively oxidized.

[0094]

[0095] (S5) A step of removing the second slag from the second molten material;

[0096] According to one embodiment of the present invention, the method for recovering valuable metal includes a step of (S5) removing second slag from the second molten material in order to remove manganese oxide and iron oxide, which are impurities, from the second molten material.

[0097] According to one embodiment of the present invention, by performing the step (S4), a second slag may be generated in the molten alloy. The second slag may include at least manganese oxide and iron oxide. For example, the second slag may include at least MnO, FeO, Fe2O3, and oxides containing manganese and iron. Manganese and iron components, particularly manganese components, correspond to impurities that lower the efficiency of a process for producing a solution containing valuable metals, which will be described later, and by removing them in step (S5), the recovery efficiency of valuable metals can be improved.

[0098] According to one embodiment of the present invention, the step of removing the second slag may include a process of removing the second slag from the surface of the second molten material and supplying oxygen to the second molten material. The second slag includes manganese oxide and iron oxide and has a lower density than the alloy, so that it exists on the surface of the second molten material. At this time, by removing the second slag existing on the surface of the second molten material, the contact area between the second molten material and oxygen is improved, so that manganese and iron remaining in the alloy can be more effectively oxidized.

[0099] According to one embodiment of the present invention, the step of removing the second slag may include a process of removing the second slag from the surface of the second molten material and supplying oxygen at least once. Specifically, the process may be repeated 1 to 10 times, 3 to 8 times, 1 to 5 times, or 5 to 10 times. By repeating the process, impurities including manganese and iron contained in the alloy can be effectively removed.

[0100] For example, the second melt may be produced by supplying oxygen to a mixture containing the alloy and the second flux for a period of 2 minutes to 30 minutes and heating the mixture. Thereafter, the heating of the second melt may be stopped, and the second slag may be removed from the surface of the second melt. Thereafter, the second melt may be heated again while supplying oxygen for a period of 2 minutes to 30 minutes, and the heating may be stopped, and the second slag may be removed from the surface of the second melt.

[0101] According to one embodiment of the present invention, in order to remove the second slag from the surface of the second molten material, methods and devices used in the art can be used without limitation. For example, when the second slag rises to the surface due to the difference in specific gravity between the second molten material and the second slag, the second slag in a molten state can be discharged outside the furnace through a tilting method or the like. In addition, the second slag can be removed from the surface of the second molten material using TRF (Tilting Rotary Furnace) equipment, TBRC (Top blown rotary converter) equipment, or the like.

[0102] According to one embodiment of the present invention, the total time for performing steps (S4) and (S5) may be 10 minutes or more and 50 minutes or less, 10 minutes or more and 30 minutes or less, 20 minutes or more and 40 minutes or less, or 20 minutes or more and 30 minutes or less. When the total time for performing steps (S4) and (S5) is within the above-mentioned range, the second flux and oxygen can be stably reacted with the alloy to effectively oxidize the manganese and iron contained in the alloy. In addition, the second slag can be effectively removed from the second molten material.

[0103]

[0104] (S6) A step of manufacturing a granulated alloy using the second molten material from which the second slag has been removed;

[0105] According to one embodiment of the present invention, the method for recovering the valuable metal comprises a step of (S6) manufacturing a granulated alloy using the second molten material from which the second slag has been removed, as a pretreatment process for recovering the valuable metal from the alloy from which impurities have been removed.

[0106] According to one embodiment of the present invention, the second molten material from which the second slag has been removed may be poured into cooling water to produce a granulated alloy. For example, a granulated alloy may be produced by dropping a predetermined amount of the second molten material from which the second slag has been removed into cooling water. Meanwhile, the granulated alloy may be produced using an air cooling method or a water cooling method, and the water cooling method may be used in terms of the efficiency of producing the granulated alloy.

[0107] According to one embodiment of the present invention, the granulated alloy may have a manganese removal rate of 99% or more compared to the alloy. Specifically, the granulated alloy may have a manganese removal rate of 99% or more and 100% or less, or 99% or more and 99.5% or less, compared to the alloy. The manganese removal rate may be calculated using the following equation 1.

[0108] [Formula 1]

[0109] Manganese removal rate (%) = (Manganese content of alloy - Manganese content of granulated alloy) / Manganese content of alloy X 100

[0110] According to one embodiment of the present invention, the granulated alloy may have a nickel content increase rate of 110% or more compared to the alloy. Specifically, the granulated alloy may have a nickel content increase rate of 110% or more and 140% or less, or 110% or more and 130% or less, compared to the alloy. The nickel content increase rate may be calculated using the following equation 2.

[0111] [Formula 2]

[0112] Nickel content increase rate (%) = (Nickel content of granulated alloy / Nickel content of alloy) X 100

[0113] According to one embodiment of the present invention, the granulated alloy may have a cobalt content increase rate of 105% or more compared to the alloy. Specifically, the granulated alloy may have a nickel content increase rate of 105% or more and 120% or less, or 110% or more and 120% or less, compared to the alloy. The cobalt content increase rate may be calculated using the following equation 3.

[0114] [Formula 3]

[0115] Cobalt content increase rate (%) = (cobalt content of granulated alloy / cobalt content of alloy) X 100

[0116] As described above, the granulated alloy has manganese removed and the contents of nickel and cobalt increased compared to the alloy before performing step (S4), so that a solution containing nickel and cobalt, which are valuable metals described below, can be effectively obtained.

[0117] According to one embodiment of the present invention, the size of the granulated alloy may be 1 mm or more. Specifically, the size of the granulated alloy may be 2 mm or more, 3 mm or more, 3 mm or more, 5 mm or more, 6 mm or more, 7 mm or more, or 10 mm or more. In addition, the size of the granulated alloy may be 20 mm or less, 15 mm or less, 10 mm or less, or 5 mm or less. By controlling the size of the granulated alloy within the above-described range, nickel and cobalt can be effectively recovered in the step of obtaining a solution containing a valuable metal described later. At this time, the size of the granulated alloy may be an average size of a plurality of granulated alloys introduced into the electrolysis tank described later.

[0118]

[0119] 1-2. Method for obtaining a solution containing valuable metals

[0120] Hereinafter, a method for recovering a valuable metal according to one embodiment of the present invention will be described, wherein a method for obtaining a solution containing the valuable metal is described. Specifically, the method for recovering the valuable metal may include a process for producing a solution containing the valuable metal by electrolyzing the granulated alloy obtained in step (S6).

[0121] (S7) A step of preparing an electrolysis tank including an electrolyte, an anode and a cathode immersed in the electrolyte, and a semi-permeable membrane provided between the anode and the cathode;

[0122] Figures 4a and 4b are drawings showing an electrolysis tank according to one embodiment of the present invention. Specifically, Figure 4a is a drawing showing an electrolysis tank (300) including an anode portion (310), a cathode portion (320), and a semi-permeable membrane (330) immersed in an electrolyte. In addition, Figure 4b is a drawing showing an anode portion (310) equipped with a receiving portion (311) capable of receiving a granulated alloy.

[0123] According to one embodiment of the present invention, the electrolysis tank can accommodate the electrolyte. The electrolyte includes an electrolyte, and the concentration of the electrolyte may be 10% or more and 30% or less. Specifically, the electrolyte may include at least sulfuric acid. In addition, the concentration of the electrolyte included in the electrolyte may be 15% or more and 25% or less, 20% or more and 25% or less, 10% or more and 25% or less, or 20% or more and 30% or less. By adjusting the concentration of the electrolyte included in the electrolyte within the above-mentioned range, valuable metals can be effectively leached from the granulated alloy.

[0124] The concentration of the electrolyte may be in wt%. That is, the concentration of the electrolyte may be 10 wt% or more and 30 wt% or less.

[0125] According to one embodiment of the present invention, the semipermeable membrane can separate the electrolyte in which the anode portion is immersed from the electrolyte in which the cathode portion is immersed. Since the semipermeable membrane has selective permeability, fine particles contained in the electrolyte can be separated based on the semipermeable membrane. This can improve the efficiency of recovering valuable metals through electrolysis.

[0126] According to one embodiment of the present invention, the thickness of the semi-permeable membrane may be 0.1 mm or more and 3 mm or less, 0.1 mm or more and 2.5 mm or less, 0.1 mm or more and 2 mm or less, 0.1 mm or more and 1.5 mm or less, 0.1 mm or more and 1 mm or less, 0.1 mm or more and 0.5 mm or less, 1 mm or more and 3 mm or less, or 2 mm or more and 3 mm or less. When the thickness of the semi-permeable membrane is within the above-mentioned range, fine particles contained in the electrolyte can be effectively separated, thereby improving the efficiency of recovering valuable metals through electrolysis.

[0127] According to one embodiment of the present invention, the semi-permeable membrane may include at least one of a cation exchange membrane and an anion exchange membrane. In addition, the semi-permeable membrane may have a current efficiency of 95% or more and 99% or less under conditions of a cell voltage of 3 V or more and 4 V or less.

[0128] According to one embodiment of the present invention, the semi-permeable membrane may include at least one of Nafion and perfluorosulfonic acid. However, the material forming the semi-permeable membrane is not limited thereto. Specifically, the semi-permeable membrane may include at least one of Nafion 211, Nafion 212, Nafion 115, Nafion 117, and Nafion 424, but the type of the semi-permeable membrane is not limited thereto. In addition, the semi-permeable membrane may include at least one of Aquivion's E98-05S, E98-09S, and E-98-15S, but the type of the semi-permeable membrane is not limited thereto.

[0129] According to one embodiment of the present invention, the anode portion may include a receiving portion capable of receiving the granulated alloy. The receiving portion may be configured to receive the granulated alloy and to allow contact between the electrolyte and the granulated alloy. By using the anode portion provided with the receiving portion, valuable metals can be stably leached from the granulated alloy.

[0130] According to one embodiment of the present invention, the materials of the positive (+) and negative (-) electrodes can be used without limitation as those used in the art. For example, each of the positive and negative electrodes can include at least one of titanium (Ti), nickel (Ni), and stainless steel.

[0131]

[0132] (S8) A step of receiving the granulated alloy in the receiving portion of the anode portion and leaching it through electrolysis;

[0133] According to one embodiment of the present invention, the valuable metal contained in the granulated alloy can be effectively leached into the electrolyte through the electrolysis. Furthermore, the copper component contained in the granulated alloy can be deposited on the cathode. Specifically, the valuable metal containing at least nickel and cobalt can be leached from the granulated alloy into the electrolyte.

[0134] According to one embodiment of the present invention, the size of the granulated alloy may be 1 mm or more. Specifically, the size of the granulated alloy may be 2 mm or more, 3 mm or more, 3 mm or more, 5 mm or more, 6 mm or more, 7 mm or more, or 10 mm or more. In addition, the size of the granulated alloy may be 20 mm or less, 15 mm or less, 10 mm or less, or 5 mm or less. When the size of the granulated alloy is within the above-mentioned range, contact between the granulated alloy accommodated in the accommodation portion and the anode portion may be facilitated. Through this, valuable metals can be effectively leached from the granulated alloy.

[0135] According to one embodiment of the present invention, the electrolysis may be performed under voltage conditions of 3.5 V or more and 15 V or less, 3.5 V or more and 10 V or less, 4 V or more and 8 V or less, 3.5 V or more and 10 V or less, or 10 V or more and 15 V or less. When the voltage conditions under which the electrolysis is performed are within the above-mentioned range, the valuable metal can be effectively leached from the granulated alloy. In addition, the electrolysis may be performed under current conditions of 7.5 A or more and 15 A or less, 10 A or more and 12.5 A or less, or 7.5 A or more and 10 A or less. By adjusting the current conditions under which the electrolysis is performed within the above-mentioned range, the valuable metal can be stably leached from the granulated alloy.

[0136] According to one embodiment of the present invention, the electrolysis can be performed under a temperature condition of 15°C or more and 80°C or less. Specifically, the temperature at which the electrolysis is performed can be 20°C or more and 70°C or less, 25°C or more and 60°C or less, 30°C or more and 45°C or less, 15°C or more and 35°C or less, 20°C or more and 30°C or less, 15°C or more and 25°C or less, 25°C or more and 35°C or less, or 40°C or more and 80°C or less.

[0137] By controlling the temperature at which the electrolytic decomposition is performed within the aforementioned range, the generation of electrolyte salts within the electrolyte can be suppressed. Specifically, by preventing the generation of sulfates within the electrolyte, the leaching efficiency through electrolysis can be improved.

[0138]

[0139] (S9) A step of filtering the extracted electrolyte to obtain a solution containing valuable metal;

[0140] According to one embodiment of the present invention, the step of obtaining a solution containing the valuable metal may include heating the leached electrolyte to a temperature of 30° C. or higher and 40° C. or lower and filtering the leached electrolyte. At this time, the leached electrolyte may refer to an electrolyte containing a valuable metal component leached from a granulated alloy by performing electrolysis. By heating the leached electrolyte to the above-mentioned temperature before filtering the leached electrolyte, the component leached from the granulated alloy can be stably dissolved. Through this, the valuable metal can be recovered more efficiently.

[0141] Any filter used in the art for filtering the above-mentioned extracted electrolyte may be used without limitation. For example, the filter may include a microfilter. Specifically, the filter may include a microfilter having a filtration size of 1 μm or more and 30 μm or less.

[0142] According to one embodiment of the present invention, the electrolyte contained in the electrolysis tank can be separated by a semi-permeable membrane. Accordingly, based on the semi-permeable membrane, a first electrolyte (+ electrolyte) in which the anode portion is immersed and a second electrolyte (- electrolyte) in which the cathode portion is immersed can be contained. At this time, the step of obtaining a solution containing the valuable metal can be performed by filtering the leached first electrolyte to obtain a solution containing the first valuable metal, and by filtering the leached second electrolyte to obtain a solution containing the second valuable metal.

[0143] According to one embodiment of the present invention, the valuable metal may include nickel and cobalt. That is, the solution containing the valuable metal may include nickel and cobalt. The solution containing the valuable metal contains a high content of nickel and cobalt, allowing for the effective recovery of nickel and cobalt from the solution containing the valuable metal.

[0144] According to one embodiment of the present invention, the solution containing the valuable metal may have a nickel content of 60% or more and 70% or less, and a cobalt content of 30% or more and 35% or less, based on the total weight of the valuable metal. That is, by performing the above-described steps, a solution having a high nickel and cobalt content can be easily obtained.

[0145]

[0146] 2. Lithium recovery method

[0147] Hereinafter, a method for recovering lithium in a method for recovering valuable metals according to an embodiment of the present invention will be described. Specifically, the method for recovering valuable metals may further include, after step (S3), a process for recovering lithium according to the following steps (S11) to (S13), in addition to the processes for recovering valuable metals described above (S4) to (S9).

[0148] According to one embodiment of the present invention, by melting the first mixture in the arc furnace, re-melting the first slag from the molten material generated, and separating the lithium solution from the fume generated in the process of melting the first slag, the recovery rate of lithium can be significantly increased, while at the same time, the effect of minimizing environmental pollution can be achieved by using a small amount of land and eco-friendly energy.

[0149]

[0150] (S11) A step of melting a second mixture including the first slag and the third flux to obtain second slag and fume;

[0151] According to one embodiment of the present invention, the method for recovering the valuable metal may include a step of (S11) melting a second mixture including the first slag and a third flux to obtain third slag and fume in order to recover lithium from the first slag at a high recovery rate.

[0152] According to one embodiment of the present invention, the step (S11) may include a step performed in a furnace different from the arc furnace. By using a furnace different from the arc furnace, energy efficiency can be improved and the lithium recovery rate can be further increased.

[0153] According to one embodiment of the present invention, various furnaces commercially available in the relevant technical field can be used without particular limitation as a means for achieving the step (S11). Specifically, the furnace may include a holding furnace. For example, the holding furnace may be an electric furnace, a plasma furnace, or a burner-type melting furnace. Specifically, the holding furnace may be any one selected from the group consisting of a direct current electric arc furnace (DC-EAF), an alternating current electric arc furnace (AC-EAF), a plasma arc furnace, and a combination thereof.

[0154] For example, the first slag may be in a molten state or a solidified state (or cooled state) before being fed into the holding furnace.

[0155] According to one embodiment of the present invention, the step (S11) may include a step of melting the second mixture at a temperature of 1400°C or higher and 1700°C or lower for 1 hour or longer and 3 hours or shorter. Specifically, the temperature of the step (S11) may be 1500°C or higher and 1650°C or lower, or 1600°C or higher and 1650°C or lower, and the melting time may be 1 hour or longer and 2 hours or shorter. By controlling the melting temperature and melting time within the above numerical ranges, the recovery rate of lithium may be further increased.

[0156] According to one embodiment of the present invention, the third flux can lower the melting point of the first slag and control the fluidity of the second mixture, while at the same time acting as a chlorine donor that induces the first slag and the third flux to react and cause LiCl(g) to smoke.

[0157] According to one embodiment of the present invention, the third flux may be any one selected from the group consisting of alkali metal chlorides, alkaline earth metal chlorides, transition metal chlorides, and mixtures thereof, and may specifically be an alkaline earth metal chloride. When an alkaline earth metal chloride is used as the second flux, the content of lithium fumed as LiCl(g) may be further increased, thereby further increasing the lithium recovery rate.

[0158] For example, the alkali metal chloride may be NaCl or KCl, the alkaline earth metal chloride may be MgCl2 or CaCl2, and the transition metal chloride may be ZnCl2.

[0159] According to one embodiment of the present invention, the content of the third flux may be 90 parts by weight or more and 130 parts by weight or less, 100 parts by weight or more and 125 parts by weight or less, 110 parts by weight or more and 125 parts by weight or less, or 115 parts by weight or more and 125 parts by weight or less, based on 100 parts by weight of the first slag. By adjusting the content of the third flux within the above numerical range, the fluidity of the second mixture is adjusted to an appropriate level, so that the recovery rate of lithium can be further increased.

[0160] In this specification, “fume” may include gas (e.g., LiCl(g)) and / or solid particles (e.g., dust) generated during the process of melting the second mixture.

[0161] According to one embodiment of the present invention, the fume may include LiCl(g). The content of lithium fumed as LiCl(g) may be 50 wt% or more, 60 wt% or more, 70 wt% or more, 80 wt% or more, or 90 wt% or more, and may be 95 wt% or less, relative to the total weight of lithium contained in the lithium-containing waste.

[0162] According to one embodiment of the present invention, the content of LiCl(g) included in the fume may be 90 vol% or more, 91 vol% or more, 92 vol% or more, 93 vol% or more, 94 vol% or more, 95 vol% or more, 96 vol% or more, 97 vol% or more, 98 vol% or more, or 99 vol% or more, and may be 99 vol% or less, based on the total volume of the fume. For example, the content of LiCl(g) may be analyzed by a gas chromatography method, a combustion ion chromatography (CIC) analysis method, and a combination thereof.

[0163]

[0164] (S12) Step of separating lithium solution from the above fume

[0165] According to one embodiment of the present invention, the method for recovering the valuable metal may include a step of separating a lithium solution from the fume (S12) to extract lithium from the fume.

[0166] According to one embodiment of the present invention, the step (S12) may include a step of dissolving the fume in water using a scrubber. For example, the fume may be dissolved in water by a nozzle (Nozzle) provided in the scrubber for spraying water, and a dissolution reaction of LiCl(g) as shown in the following reaction formula 1 may proceed. For example, the lithium solution may include LiCl(aq).

[0167] [Reaction Formula 1]

[0168] LiCl(g)+H2O(l)->Li + (aq)+Cl - (aq)

[0169] According to one embodiment of the present invention, the scrubber may include one or more wet scrubbers. For example, the wet scrubber is not particularly limited, but may include at least one selected from the group consisting of a packed type scrubber, a spray type scrubber, a cyclone scrubber, a venturi scrubber, and an ejector scrubber.

[0170] According to one embodiment of the present invention, the scrubber may include two or more interconnected sub-scrubbers. Specifically, when two or more sub-scrubbers are used, a dust collector can be omitted, reducing manufacturing equipment costs and thus improving the economic efficiency of the process.

[0171] Hereinafter, with reference to FIG. 3, a scrubber used in a process for recovering lithium in a method for recovering valuable metals according to one embodiment of the present invention will be described.

[0172] Figure 3 is a schematic diagram of a wet scrubber used in a method for recovering valuable metals according to one embodiment of the present invention.

[0173] Referring to FIG. 3, the wet scrubber (200) may include a first sub-scrubber (200a), a second sub-scrubber (200b), and a connecting portion (200c) connecting the first and second sub-scrubbers (200a, 200b) to each other.

[0174] According to one embodiment of the present invention, the first sub-scrubber (200a) may include a first gas inlet (F), a first charge (230a), a first water supply pipe (210a), a first nozzle (210n1), and a first recovery unit (220a).

[0175] According to one embodiment of the present invention, the first gas inlet (F) may be an inlet through which the fume is injected. For example, the fume generated in the second melting process may be introduced into the internal space of the first sub-scrubber (200a) through the first gas inlet (F) and transferred to the first filler (230a) described below.

[0176] According to one embodiment of the present invention, the first charge (230a) may be a region in which a LiCl(aq) and residual gas (R_G) generation reaction occurs by further increasing the contact area between the fume introduced through the first gas inlet (F) and the water sprayed from the first nozzle (210n1) described later. For example, the first charge (230a) may be in the form of a layer or a form in which particles are clustered. For example, the first charge (230a) may include at least one selected from the group consisting of ceramic particles, metal materials, and polyolefins. In one example, the metal material may include stainless steel, and the polyolefin may include polypropylene.

[0177] For example, the reaction of the above reaction formula 1 may proceed on the surface of the first charge (230a). As the above reaction formula 1 proceeds on the first charge (230a), LiCl(aq) that is primarily recovered by the first recovery unit (220a) to be described later; and a residual gas (R_G) containing lithium may be generated. At this time, the residual gas (R_G) may be transferred to the second sub-scrubber (200b) through the connection unit (200c) to be described later.

[0178] According to one embodiment of the present invention, the first water supply pipe (210a) can supply water to the first nozzle (210n1) described later via a water pump (not shown). For example, the first water supply pipe (210a) can be placed on the first charging material (230a).

[0179] According to one embodiment of the present invention, the first nozzle (210n1) can spray water supplied from the first water supply pipe (210a) onto the first charging material (230a). Through this, a reaction can occur on the surface of the first charging material (230a) in which LiCl(g) reacts with water to produce LiCl(aq). For example, the number of the first nozzles (210n1) is not limited to the embodiment illustrated in FIG. 3 and can be modified in various ways.

[0180] According to one embodiment of the present invention, the first recovery unit (220a) can primarily recover LiCl(aq) generated from the first charge (230a). For example, the first recovery unit (220a) can be arranged below the first sub-scrubber (200a). The primarily recovered LiCl(aq) can be discharged through the first outlet (L1). For example, the first outlet (L1) can be connected to one side of the first recovery unit (220a).

[0181] According to one embodiment of the present invention, the second sub-scrubber (200b) may include a second charging material (230b), a second water supply pipe (210b) for supplying water to a second nozzle (210n2) to be described later, a second nozzle (210n2), and a second recovery unit (220b). The description of the second sub-scrubber overlaps with the description of the first sub-scrubber, and is therefore briefly described or omitted.

[0182] For example, the residual gas (R_G) transferred into the interior of the second sub-scrubber (200b) through the above-mentioned connection portion (200c) can react with water sprayed through the second nozzle (210n2) on the surface of the second charge (230b). Accordingly, the secondarily recovered LiCl(aq) can be recovered through the second recovery portion (220b) and, specifically, can be discharged through the second outlet (L2).

[0183] According to one embodiment of the present invention, the first filler (230a) may be the same as or different from the second filler (230b).

[0184] According to one embodiment of the present invention, the wet scrubber (200) may additionally be equipped with equipment for combining lithium solutions discharged through the first and second outlets (L1, L2).

[0185] According to one embodiment of the present invention, the recovery solution recovered from the first and second recovery units (220a, 220b) may contain impurities in the form of fine particles dissolved in water in addition to the lithium solution.

[0186]

[0187] (S13) Step of concentrating the lithium solution

[0188] According to one embodiment of the present invention, the method for recovering valuable metals may further include a step of concentrating the lithium solution (S13) to recover LiCl containing a high concentration of Li ions. Through this, a high-concentration lithium solution can be obtained.

[0189] According to one embodiment of the present invention, the concentration of lithium ions in the high-concentration lithium solution may be 10,000 ppm or more, 15,000 ppm or more, or 20,000 ppm or more, and 30,000 ppm or less, or 25,000 ppm or less. Here, the concentration of lithium ions may be analyzed by an ICP-OES (Inductively coupled plasma optical emission spectroscopy) method.

[0190] According to one embodiment of the present invention, the concentration of chloride ions in the high-concentration lithium solution may be 170,000 ppm or more and 200,000 ppm or less. Here, the concentration of chloride ions may be measured according to a combustion ion chromatography (CIC) analysis method.

[0191] For example, a common concentrator commercially available in the relevant technical field may be used as a means for concentrating the lithium solution. Specifically, the concentrator may be a centrifugal vacuum concentrator, a rotary vacuum concentrator, a rotary evaporator, or a centrifugal vacuum concentrator.

[0192] According to one embodiment of the present invention, as a method for concentrating the lithium solution, a method of using at least one type of waste heat selected from the group consisting of heat energy generated during the firing process, heat energy generated during the melting process of the arc furnace, heat energy generated during the melting process of the second mixture, heat energy of the first slag, heat energy of the alloy, and heat energy of the second slag may be used. By using the waste heat, carbon dioxide emitted during the process can be reduced while simultaneously saving electric energy, thereby lowering the manufacturing cost. For example, in order to use the waste heat, a duct capable of transporting the waste heat can be connected to at least one of each of the devices for achieving steps (S1) to (S3), and (S11).

[0193] According to one embodiment of the present invention, a method of concentrating the lithium solution may be a method utilizing any one electric energy selected from the group consisting of an arc furnace, a holding furnace, and a combination thereof. By utilizing the electric energy, carbon dioxide emissions during the process can be reduced, while electric energy consumption can be reduced, thereby lowering manufacturing costs.

[0194] Figure 5 is a flowchart showing a method for recovering lithium in a method for recovering valuable metals according to one embodiment of the present invention.

[0195] Referring to FIG. 5, a method for recovering lithium according to an embodiment of the present invention may include the steps of (a) preparing black mass by crushing waste lithium-ion batteries; (b) calcining the black mass; (c) mixing the calcined black mass with a first flux to prepare a first mixture; (d) first melting the first mixture in a direct current electric arc furnace under reducing conditions to obtain a melt; (e) tapping the melt to separate the first slag and the alloy; (f) mixing and second melting the first slag with a second flux in a holding furnace; (g) bringing fumes generated in the second melting process into contact with a scrubber to prepare a lithium solution; and (h) concentrating the lithium solution.

[0196] According to one embodiment of the present invention, the recovery rate of lithium recovered by the lithium recovery method may be 90% or more, 91% or more, 92% or more, or 93% or more, and may be 99% or less, or 95% or less. For example, the lithium recovery rate may be the ratio of the lithium ion content in the concentrated lithium solution to the lithium content in the black mass.

[0197]

[0198] 3. Lithium compounds

[0199] According to one embodiment of the present invention, a lithium compound recovered through a method for recovering lithium according to the above-described embodiment is provided.

[0200] In this specification, "lithium compound" may be defined as a compound containing lithium ions, which is a result recovered by the method for recovering lithium or a product using the same. For example, the result recovered by the method for recovering lithium may include LiCl(l), and the product may be lithium carbonate (Li2CO3) produced by the reaction of lithium ions with carbonate or carbon dioxide. For example, the lithium carbonate may be widely applied in various fields such as electrode materials for lithium ion batteries, electrolyte materials, glass production, and aluminum production.

[0201] According to one embodiment of the present invention, the purity of the lithium compound may be 90% or more, 91% or more, 92% or more, 93% or more, 94% or more, 95% or more, 96% or more, 97% or more, 98% or more, or 99% or more, and may be 99.5% or less.

[0202]

[0203] Hereinafter, embodiments of the present invention will be described in detail so that those skilled in the art can easily implement the present invention. However, this is merely an example, and the scope of the present invention is not limited by the following contents.

[0204] As used herein, terms such as "about" or "substantially" mean a reasonable amount of variation from the term that does not significantly alter the final result. These terms may be interpreted to include a variation of at least ±5% or at least ±10%, provided that such variation does not alter the meaning of the term and render it invalid.

[0205]

[0206] Example

[0207] Steps to prepare black mass by crushing waste lithium-ion batteries:

[0208] Approximately 100 kg of waste lithium-ion batteries were crushed to approximately 100 mesh or less to prepare black mass.

[0209]

[0210] Steps for calcining the above black mass:

[0211] In an air atmosphere, the black mass was heated from room temperature to about 750°C at a heating rate of about 100°C, and then calcined at about 750°C for 12 hours to remove carbonized substances. The resultant product from which the carbonized substances were removed was sufficiently cooled at room temperature for 240 minutes, thereby producing calcined black mass having the composition shown in Table 1 below. The composition of the calcined black mass was measured using an ICP-OES (Inductively coupled plasma optical emission spectroscopy) device from Perkin Elmer.

[0212] ClassificationNiCoMnLippm295,250128,900102,30064,785

[0213] A step of preparing a first mixture by mixing the above-mentioned calcined black mass and the first flux:

[0214] A first mixture was prepared by mixing about 60 wt% of the calcined black mass, about 7 wt% of Al2O3, about 17 wt% of CaO, and about 16 wt% of SiO2 using a drum mixer for about 1 hour.

[0215]

[0216] A step of melting the first mixture in a direct current electric arc furnace under reducing conditions:

[0217] Before the first mixture was introduced, a batch type DC-Electric Arc Furnace (DC-EAF) was preheated to about 1,500°C for about 2 hours. At this time, the output during preheating was about 80 to 100 kW and the input voltage was 130 to 150 V.

[0218] Meanwhile, the direct current electric arc furnace includes an electrode rod equipped with a carbon electrode. Since the carbon electrode reduces the calcined black mass, it is possible to easily produce an alloy, which will be described later, by reducing the valuable metal contained in the calcined black mass without adding a separate reducing agent.

[0219] After the first mixture was introduced through a hopper equipped in the above-described preheated direct current electric arc furnace, it was melted at 1,500°C for 3 hours under an air atmosphere.

[0220]

[0221] Step of tapping the first molten material to separate the first slag and alloy:

[0222] The first molten material was tapped to separate the first slag and the alloy. At this time, the composition of the first slag is as shown in Table 2 below, and the composition of the alloy is as shown in Table 3 below.

[0223] Meanwhile, the compositions described in Tables 2 and 3 below were analyzed using the ICP-OES (Inductively coupled plasma optical emission spectroscopy) method.

[0224] ClassificationNiCoMnLiAlCuSiFeCaMgSUMppm0089,92068,63094,3500124,0000176,6008,180561,680wt%0016.012.216.8022.1031.41.5100

[0225] ClassificationNiCoMnLiAlCuSiFePCaCrSUMppm625,300334,60092,690058361,8003,54927,57011,9407151,9211,160,668wt%53.828.88.000.15.30.32.41.00.10.2100

[0226] A step of producing a second molten material using an alloy and a second flux, and removing a second slag;

[0227] SiO2 was prepared as a second flux. The alloy and second flux separated above were introduced into a TBRC furnace. At this time, about 7 parts by weight of the second flux was introduced per 100 parts by weight of the alloy.

[0228] Afterwards, oxygen was added to the TBRC at 9 kgf / cm 2 The TBRC furnace was heated to a temperature of approximately 1,400°C for 5 minutes under partial pressure conditions. Through this, the second molten material was produced. The oxygen supply and heating were stopped, and the second slag that floated on the surface of the second molten material was removed. Afterwards, oxygen was supplied to the TBRC furnace at 9 kgf / cm 2 After supplying oxygen under partial pressure conditions for 5 minutes and heating the TBRC furnace to a temperature of approximately 1,400°C, the oxygen supply and heating were stopped, and the process of removing the second slag that floated on the surface of the second molten material was repeated three more times. Finally, heating was performed for a total of 20 minutes, and the second slag was removed three times.

[0229] FIG. 6 is a photograph of a second molten substance produced in an embodiment of the present invention and second slag removed from the second molten substance. Specifically, FIG. 6 (a) is a photograph taken during the process of producing the second molten substance in the embodiment, and FIG. 6 (b) is a photograph of second slag removed from the second molten substance.

[0230]

[0231] Preparation of granulated alloys;

[0232] The second molten material from which the second slag had been removed was dropped into cooling water to obtain multiple granulated alloys having a size of approximately 4 to 7 mm. Granulated alloys having a size of less than 4 mm were remelted and reused. The composition of the granulated alloys was analyzed using ICP-OES (Inductively coupled plasma optical emission spectroscopy), and the results are shown in Table 4 below.

[0233] ClassificationNiCoMnLiAlCuSiFePCaCrSUMppm776,400375,3500098167,4302,8128303,81769101,228,311wt%63.2130.56000.085.490.230.070.310.060100

[0234] Referring to Tables 3 and 4 above, it was confirmed that the granulated alloy had a manganese removal rate of 100%, a nickel content increase rate of approximately 124%, and a cobalt content increase rate of approximately 112% compared to the alloy.

[0235]

[0236] Steps for preparing an electrolysis tank;

[0237] As shown in Figures 4a and 4b, an electrolysis tank was prepared.

[0238] An electrolyte solution with a sulfuric acid concentration of 25% was prepared. As a semi-permeable membrane, N-424 (Duran) with a current efficiency of 97% and a thickness of 0.18 mm was prepared under the conditions of a cell voltage of 3.74 V and 2% NaOH. Afterwards, the electrolyte solution (25 ℃) was introduced into each electrolysis tank separated by a semi-permeable membrane (N-424, Duran).

[0239] Afterwards, the granulated alloys were placed in the receiving portion of the anode, the anode was immersed in the first electrolyte (+ electrolyte), and the cathode was immersed in the second electrolyte (- electrolyte).

[0240]

[0241] Step of leaching through electrolysis;

[0242] Electrolysis was performed for 3 days under conditions of 15 A current and 4 V voltage to leach valuable metals.

[0243]

[0244] A step of obtaining a solution containing a valuable metal;

[0245] The first electrolyte and the second electrolyte, which had undergone electrolysis, were each heated to 40°C and then filtered using a microfilter (Hyundai Microfilter No. 53; filtration size: 1 to 2 μm) to produce a solution containing valuable metals.

[0246] The components of the first solution obtained from the first leached electrolyte (leached + electrolyte) were analyzed using the ICP-OES method, and the results are shown in Table 5 below. In addition, the components of the second solution obtained from the second leached electrolyte (leached - electrolyte) were analyzed using the ICP-OES method, and the results are shown in Table 6 below.

[0247] CategoryNiCoMnLiAlCuSiFePCaCrSUMPpm88,98850,43800765,54840020145,056wt%61.34734.771000.0523.8250.003000.0010100

[0248] ClassificationNiCoMnLiAlCuSiFePCaCrSUMppm37,13819,9710057150001057,182wt%64.94734.925000.1000.0260000.0020100

[0249] Referring to Tables 5 and 6 above, it was confirmed that the content of nickel contained in the solution containing valuable metals was 60% or more, and the content of cobalt was 34% or more.

[0250]

[0251] Meanwhile, in order to obtain a lithium solution from the first slag separated from the first melt, the following steps were performed.

[0252]

[0253] Step of mixing and melting the first slag with the third flux in a holding furnace:

[0254] A first slag separated from the first molten material was prepared. Thereafter, the first slag was placed in a holding furnace different from the DC electric arc furnace and melted at approximately 1,600°C for 0.5 hours. During the melting of the first slag, a third flux composed of 100 parts by weight of CaCl2 and 25 parts by weight of SiO2 was placed relative to 100 parts by weight of the first slag, thereby performing a melting step. As a result, a second slag having the composition shown in Table 7 below and fume were obtained. At this time, the second slag was analyzed by the ICP-OES (Inductively coupled plasma optical emission spectroscopy) method.

[0255] Meanwhile, LiCl (g) contained in the fume was analyzed through a LiCl aqueous solution sample prepared by dissolving the fume in 1 ton of water. Specifically, Li in the sample + It was analyzed to be approximately 1,532 ppm using the ICP-OES (Inductively coupled plasma optical emission spectroscopy) method.

[0256] ClassificationNiCoMnLiAlCuSiFeCaMgSUMppm0031,910083,1300108,4000330,1007,308560,848wt%006%015%019%059%1%100%

[0257] Step of manufacturing a lithium solution by contacting the fume generated in the above melting process with a scrubber:

[0258] The above fume was passed through a two-stage wet scrubber to dissolve Li in water. The fume generated during the melting process was recovered by suction, and lithium was recovered in the first sub-scrubber. The residual gas that passed through without being captured during the process was recovered by suction in the second sub-scrubber to obtain a lithium solution.

[0259]

[0260] Step of concentrating the above lithium solution to recover lithium:

[0261] The lithium solution was concentrated and evaporated at 90°C for 5 hours using an evaporator to obtain a high-concentration lithium solution. At this time, the concentration of lithium ions in the high-concentration lithium solution was 20,000 ppm or more based on the ICP-OES (Inductively coupled plasma optical emission spectroscopy) method. In addition, Cl in the high-concentration lithium solution - It was measured to be approximately 172,399 ppm according to the Combustion Ion chromatography (CIC) analysis method.

[0262]

[0263] [Experimental Example: Measurement of Lithium Recovery Rate and Recovery Time]

[0264] The recovery rate of lithium was calculated according to Equation 4 below.

[0265] [Formula 4]

[0266] Lithium recovery rate (%) = (O L / I L ) x 100

[0267] In the above equation 4, I L is the lithium content in black mass (kg), and O L is the content (kg) of lithium ions in the recovered concentrated lithium solution.

[0268] Example 1 of the first mixture melting method Direct current electric arc Lithium recovery rate (%) 93% Lithium recovery time (unit: hour, time) 24

[0269] Referring to Table 8 above, it was confirmed that when lithium was recovered using the method according to Example 1, the lithium recovery rate was 93% or more, and the lithium recovery time was significantly shortened to 24 hours.

[0270]

[0271] Comparative Example 1

[0272] In the above example, a second melt was prepared in the same manner as in the example, except that only the second flux was supplied to the alloy without supplying oxygen during the preparation of the second melt, and after removing the second slag, a granulated alloy was obtained.

[0273] Thereafter, the manganese removal rate of the granulated alloy was measured using the same method as in the example, and the measurement result confirmed that the manganese removal rate was 41%.

[0274]

[0275] Comparative Example 2

[0276] In the above example, a second melt was prepared in the same manner as in the example, except that only oxygen was supplied to the alloy without adding a second flux when preparing the second melt, and after removing the second slag, a granulated alloy was obtained.

[0277] Thereafter, the manganese removal rate of the granulated alloy was measured using the same method as in the example, and the measurement result confirmed that the manganese removal rate was 53%.

[0278]

[0279] Reference Example 1

[0280] Among the granulated alloys manufactured in the above examples, multiple granulated alloys having a size of approximately 1 mm were obtained. Thereafter, a solution containing a valuable metal was manufactured from the granulated alloy using the same method as in the examples.

[0281] The second solution obtained from the leached second electrolyte (leached - electrolyte) was analyzed for components using the ICP-OES method. The nickel content in the second solution was 29,880 ppm, and the cobalt content was 14,168 ppm.

[0282] In the second solution obtained from the second electrolyte (leached - electrolyte) leached in the above example, the nickel content was 37,138 ppm, and the cobalt content was 19,971 ppm. Referring to this, it can be seen that even when using the same amount of granulated alloy, Reference Example 1, in which the size of the granulated alloy was 1 mm, has a smaller content of recovered nickel and cobalt compared to the examples.

[0283]

[0284] In the second solution obtained from the second electrolyte (-electrolyte) in the above example, the nickel content was 37,138 ppm, and the cobalt content was 19,971 ppm. Referring to this, even when using the same amount of granulated alloy, it can be seen that Comparative Example 3, in which the size of the granulated alloy is less than 5 mm, has a lower content of recovered nickel and cobalt compared to the example.

[0285]

[0286] The features described in the above-described embodiment may be combined with other embodiments unless explicitly stated otherwise. Furthermore, while the preferred embodiments of the present invention have been described in detail above, the scope of the present invention is not limited thereto. Various modifications and improvements made by those skilled in the art utilizing the basic concepts of the present invention defined in the following claims also fall within the scope of the present invention.

[0287] [Explanation of symbols]

[0288] 100: Direct current electric arc furnace

[0289] 10: Electrode rod

[0290] 20: Lower electrode

[0291] 30: Separation section

[0292] 30a: 1st separation section

[0293] 30b: Second Separation Section

[0294] 40: Exhaust

[0295] 50: Cooling section

[0296] 50a: 1st cooling section

[0297] 50b: Second cooling section

[0298] 200: Wet scrubber

[0299] 200a: 1st sub-scrubber

[0300] F: 1st gas inlet

[0301] 230a: First charge

[0302] 210a: First water supply pipe

[0303] 210n1: Nozzle 1

[0304] 220a: 1st Recovery Unit

[0305] L1: Exit 1

[0306] 200b: Second sub-scrubber

[0307] 230b: Second charge

[0308] 210n2: Second nozzle

[0309] 220b: Second Recovery Unit

[0310] L2: Exit 2

[0311] 200c: Connection

[0312] R_G: Residual gas

[0313] 300: Electrolysis tank

[0314] 310: Bipolar

[0315] 311: Reception

[0316] 320: Cathode

[0317] 330: Semipermeable membrane

Claims

A step of preparing a first mixture comprising a powder obtained by crushing lithium-containing waste and a first flux; A step of melting the first mixture in an arc furnace under reducing conditions to produce a first melt; A step of separating the first slag and alloy from the first molten material; A step of removing impurities from the above alloy and producing a granulated alloy; A step of preparing an electrolysis tank including an electrolyte, an anode and cathode immersed in the electrolyte, and a semi-permeable membrane provided between the anode and cathode; A step of receiving the granulated alloy in the receiving portion of the anode portion and leaching it through electrolysis; and A method for recovering valuable metals, comprising: a step of filtering the extracted electrolyte to obtain a solution containing valuable metals. In the first paragraph, The above lithium-containing waste, A method for recovering valuable metals, comprising at least one selected from the group consisting of waste lithium-ion batteries, waste electrodes, and waste cathode active materials. In the first paragraph, The above first flux is Al2O 3, CaO, SiO2, and A method for recovering a valuable metal, comprising at least one selected from the group consisting of Fe2O3. In the first paragraph, In the above arc, A method for recovering valuable metals, wherein the method comprises any one selected from the group consisting of a direct current electric arc furnace, an alternating current electric arc furnace, a plasma arc furnace, and combinations thereof. In the first paragraph, The step of manufacturing the first melt is as follows: A method for recovering valuable metals, which is performed by a direct current electric arc furnace under an output condition of 50 kW or more and 3,000 kW or less. In the first paragraph, A method for recovering valuable metals, wherein the size of the granulated alloy is 1 mm or more. In the first paragraph, A method for recovering precious metals, wherein the above electrolysis is performed under voltage conditions of 3.5 V or more and 15 V or less. In the first paragraph, A method for recovering valuable metals, wherein the above electrolysis is performed under temperature conditions of 15°C or higher and 80°C or lower. In the first paragraph, A method for recovering valuable metals, wherein the electrolyte solution contains an electrolyte, and the concentration of the electrolyte is 10% or more and 30% or less. In the first paragraph, The step of obtaining a solution containing the above valuable metal is: A method for recovering valuable metals, comprising heating the extracted electrolyte to a temperature of 30°C or higher and 40°C or lower and filtering the same. In the first paragraph, A step of melting a second mixture containing the first slag and the third flux to obtain a third slag and fume; and A method for recovering valuable metals, further comprising a step of separating a lithium solution from the above fume.

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