Recycling of metallurgical residues

Abstract

Metal recycling comprising melting in an electric electrode furnace tank (3, 4) a material to be treated comprising metal residues to which lime or a source of lime is added, pouring a slag (12) formed during the melting of said material to be treated through a liquid discharge path (5) to the outside of the tank and recovering from the tank a produced ferrometallic alloy (13). The furnace comprises two electrodes (3, 4) placed above the molten material and supplied with direct current, an opening in the tank of the liquid discharge path (5) and the two electrodes (3, 4) being placed in a common vertical plane. Figure: 1

Description

Title of the invention: Recycling of metallurgical residues

[0001] The invention falls within the field of processes for recycling electrochemical metallurgical residues, such as batteries or accumulators, or petrochemical residues, such as catalysts.

[0002] We are interested in residues containing iron, nickel, molybdenum, chromium, tungsten, cobalt, manganese, copper, zinc or lithium. It is possible and known to recover these elements in the form of a ferroalloy.

[0003] Without seeking to be exhaustive, the residues concerned are identified as follows, which are co-products, generated in significant quantities, of industrial processes producing main products constantly demanded by the modern economy: used chemical and petrochemical catalysts, containing the metals nickel, molybdenum, cobalt, tungsten, vanadium, copper in the form of oxides or sulfides or chlorides; used portable saline or alkaline or zinc-air batteries, mainly containing the metals iron, zinc and manganese; dust and sludge from steelmaking or steel surface treatment, generally containing the metals iron, nickel, chromium, molybdenum, copper, tungsten, cobalt and zinc; used portable and electrified vehicle accumulators or new production waste containing the metals iron, nickel, cobalt, manganese, copper, aluminum and lithium.

[0004] The recycling of the residues thus cited is known mainly in the form of pyrometallurgical processes, that is to say using heat to separate the metals.

[0005] We can thus cite: the recycling by the Waelz process of dust and sludge from steelworks involved in the manufacture of unalloyed steels, rich in iron and zinc - the Waelz process nevertheless only recovers zinc; the more complete recycling of these dusts and sludges in the electric arc furnace supplied with three-phase current, comprising three graphite electrodes, possibly preceded by a preheating-pre-reduction stage, and with the use of an electric furnace using plasma torches where appropriate; the recycling again of dust and sludge from steelworks involved this time in the manufacture of alloyed and stainless steels, rich in nickel, chromium, manganese and zinc, in the free electric arc furnace supplied with three-phase current, comprising 3 graphite electrodes. Three-phase current is also used most often with a submerged arc furnace.It is also possible to carry out a two-stage treatment, namely a preliminary reduction in a rotating hearth or multiple hearth furnace, then melting in a three-phase electric submerged arc furnace; the recycling of alkaline and saline batteries, rich in Iron, Manganese and Zinc by crushing separating the steel cans, then treatment of the remaining mass in the furnace. Waelz - but this only recovers Zinc, as has already been said; the direct treatment of these same batteries in a three-phase electric furnace, free or submerged arc; or the recycling of petrochemical catalysts of the Iron-Nickel-Molybdenum type in a three-phase electric furnace, free or submerged arc.

[0006] A comparison of these different solutions shows that the three-phase free arc electric furnace offers a high-performance solution. This is observed by the ratio of the specific melting power, namely the active melting power, to the surface area of ​​the metal bath. This ratio is 1 to 2 MW / m2 for the three-phase free arc electric furnace compared to 0.25 to 0.4 MW / m2 in the case of a three-phase submerged arc furnace.

[0007] However, the three-phase free arc electric furnace has a productivity limitation caused by the inhomogeneity and the difficulty of cleaning the slag, that is to say the low density co-product of the separation reaction, when the material to be recycled is rich in mineral gangue or when its treatment requires significant inputs of mineral materials such as lime or magnesia, which is the case for the catalysts and batteries and accumulators in question, and gives rise to a significant production of slag, from 0.2 tonnes to more than 1 tonne of slag per tonne of material to be treated. The used catalysts themselves contain a significant base (60% or more) of alumina, which must be compensated by an equivalent input of lime; Alkaline and saline batteries contain little gangue, but the slag formed contains a lot of manganese oxide MnO, which requires the addition of lime, magnesia, possibly alumina and silica for further recovery.

[0008] In addition, in all cases, foam is produced due to the formation of gases - carbon monoxide CO or zinc metal in the form of vapor, or even sulfur oxides SOx - by the reduction of the metallic chemical species present. These large and voluminous quantities of slag require frequent shutdowns to clean the furnace clogged with slag.

[0009] Furthermore, the slag forms accretions which block the pouring holes of the metal and the slag, even when the slag is poured through a dedicated door. These accretions freeze the slag at its free surface due to lack of heating, thus limiting the height of slag in the furnace and therefore the duration of the melting cycle.

[0010] The mineral components contained in the materials to be treated, catalysts or batteries, or added for the latter, are mainly constituted by alumina, silica, lime, and magnesia. To the mineral gangue contained (especially in the catalysts) are added additional materials such as lime and / or limestone (a limestone source of lime), and / or dolomite (a carbonate rock source of magnesia) in order to obtain a ratio of mass of lime plus mass of magnesia to mass of alumina plus mass of silica strictly greater than 1.2 in order to avoid or limit the chemical corrosion of the refractory elements of the wall of the furnace. These additions further increase the mineral fraction of the mixture which ultimately ends up in the quantity of slag formed and accentuate the difficulty in keeping it fluid and separating it from the recoverable metal alloy.

[0011] In current practice, this leads to wasted time, or excessive energy consumption for the purpose of overheating to keep the slag liquid, or even to the addition of a compound with a fluxing effect such as fluorspar or sodium compounds to melt the accretions. However, these additions are aggressive towards the refractory components of the furnace wall - hence in all cases, significant losses in productivity and energy. In addition, the practice of additions is detrimental to the service life of the refractory elements of the furnace wall, and causes significant production stoppages for their repair or replacement.

[0012] Furthermore, if the accretions are not controlled, they can cause electrode breakages which are particularly detrimental to the operation of the furnaces and the operating cost.

[0013] Furthermore, document EP1983066_Al discloses a method for treating spent catalysts, using a cast iron bath base, adding lime, and stirring by injecting gas into an electric arc furnace for melting the catalysts. Iron oxide and gaseous oxygen allow the phosphorus content in the reaction product to be reduced. To promote the reaction, limestone and magnesia, which are slag-forming products, are added, and the slag is poured several times in an intermediate manner before the alloy is poured. The gas injection is particularly vigorous to allow the melting of the catalysts. The bath base can be at 1550°C and the furnace can be a 3MVA free arc furnace.This process, although interesting, does not solve the difficulty that arises when the large quantity of slag leads to overheating and consuming too much energy, or to interrupting the process very often to pour off the slag.

[0014] The invention therefore aims to enable the above-mentioned metallurgical residues to be treated in a free arc electric furnace, without the quantities of slag produced constituting a hindrance to the process. Generally speaking, the subject of interest is metal residues associated with oxides which are little or very little reducible, giving rise to the formation of a slag, in particular aluminium oxide or manganese oxide, and which may require additions to be fusible.

[0015] This problem is overcome by means of an electric furnace installation and method defined as follows: the metal recycling method comprises melting in an electric electrode furnace tank a material to be treated comprising metal residues and poorly reducible oxides to which lime or a source of lime is added, and optionally iron oxide is also added, casting via a liquid discharge route to the outside of the tank of a slag formed during the melting and a recovery outside the tank of a ferrometallic alloy also produced during the melting. These are known principles, which make it possible to separate on the one hand the metals whose oxides are reducible, and which are isolated in the form of a ferroalloy, and on the other hand the metals whose oxides are much less reducible, which remain in a slag in the form of oxide, the slag being possibly, if necessary, made fusible by the addition of lime or limestone.

[0016] Remarkably, the method comprises the supply of direct current by two furnace electrodes placed above the molten material. And to take advantage of this direct current supply which only mobilizes two electrodes, an opening in the tank for the evacuation of liquid for said casting, and the ends of the two electrodes for the direct current supply are placed in a common vertical plane.

[0017] Thanks to these characteristics, which define particular and original arrangements for the method of supplying electrical energy, the loading of materials and the metal and slag outlets, and the position of the arcs, simplified management of the liquid bath is obtained. In fact, a fluid slag is kept in front of these pouring orifices, and thus the operations of pouring the metal and slag are made easier and faster. It is then possible to minimize the addition of flux, which is detrimental to the life of the refractory elements of the furnace walls, and to maintain the energy inputs at a moderate level, by heating less or for a shorter time.

[0018] The process is applied to the recycling of used catalysts or batteries, to recover molybdenum, cobalt, nickel, chromium, tungsten, or manganese in the form of ferroalloy, and separate these from a slag of lime, silica, alumina and possibly manganese oxide (which in this case can be usefully recovered, in particular if the manganese oxide is mixed with lime), in the case of batteries. The process is of particular interest when the slag is particularly abundant, and makes it possible to minimize the electrical energy used as much as possible by heating only to the strict minimum, that is to say slightly above the melting temperature of the recovered alloy, therefore above 1300° and in any case below 1600°.

[0019] According to advantageous and optional characteristics:

[0020] A drop point in the molten material of the material to be treated can be placed in the common vertical plane of the opening in the tank of the liquid discharge path and the ends; A DC2 type furnace (i.e., direct current with 2 graphite electrodes) of 2.5 m in diameter is envisaged, with a distance between the electrode closest to the slag tap hole, and the slag tap hole of approximately 70 cm. Overall, it is desired that the two electrodes, the drop point of the material to be treated, the slag tap hole and the metal discharge channel are placed in a plane. It is specified that the material to be recycled does not fall vertically and that therefore it is not its point of introduction but its point of fall which must be in the indicated plane. The fact of having the two electrodes, the point of fall and the slag tap hole, or even the metal discharge channel in the same plane makes it possible to direct the heat from the arcs created by the electrodes into the areas which need to be heated as a priority, namely the entrance to the slag tap hole, the mixing zone at the point of fall, and the entrance to the metal discharge channel.

[0021] The method may further comprise stirring the molten material by injecting gas from a bottom wall of the tank.

[0022] This limits the accretions of materials to the wall of the bath and greatly limits electrode breakages.

[0023] The method may further comprise a progressive loading of the material to be treated into the tank and several successive evacuations of the slag, separated by melting phases, before an evacuation of the ferrometallic alloy produced.

[0024] The method may further comprise capturing volatile dust escaping above the molten material and reintroducing it into the tank during melting.

[0025] Slag pouring and recovery from the alloy tank may be carried out by tilting the tank respectively to a first side and a second side opposite the first side.

[0026] The tank may be a rack-and-pinion cradle tank operated by a jack.

[0027] After the recovery of the produced ferrometallic alloy from the tank, a base of the produced ferrometallic alloy bath can be kept in the tank and a material to be treated comprising metallic residues is again introduced to continue the treatment.

[0028] The material to be treated may consist of raw used catalysts and is added, for fusion, to mineral materials, including lime. It may also consist of used batteries or accumulators.

[0029] Before the recovery from the tank of ferrometallic alloy produced and after a pouring of the slag formed during the melting, a treatment of the molten material by chemical desulfurization or dephosphorization agents can be carried out in the tank.

[0030] The material to be treated may comprise at least nickel, molybdenum, iron or zinc.

[0031] Lime, limestone or dolomite may be added to the material to be treated in order to obtain in the molten material a ratio of mass of lime plus mass of magnesia to mass of alumina plus mass of silica strictly greater than 1.2.

[0032] The invention will be better understood, and other aims, characteristics, details and advantages thereof will appear more clearly in the explanatory description which follows, made with reference to the appended drawings given solely by way of example. illustrating several embodiments of the invention and in which:

[0033] [Fig.l] is a schematic view of an oven according to the invention, seen from the side.

[0034] [Fig.2] is a diagram showing the steps of the invention.

[0035] [Fig. 3] is a schematic view of an oven according to the invention, seen from above.

[0036] [Fig.l] The installation comprises an electric furnace 1, having a charging opening 2 in the vault, two vertical electrodes 3 and 4 mounted in the vault, a pouring door 5 intended for the slag, a pouring channel 6 intended for the metal alloy produced, and two stirring gas inlet pipes 9 at the bottom of the hearth. The furnace 1 can in particular be equipped with two stirring gas injection orifices through the hearth of the furnace.

[0037] The bottom or lower wall of the furnace 7 is referenced in the figure, as well as the stirring nitrogen 8 coming out of the stirring gas inlet pipes 9, this gas stirring the bath of material and in particular the slag 12 and the metal 13 which separate from each other during melting. The nitrogen can be replaced by argon, or another neutral gas or a neutral or slightly reactive gas mixture.

[0038] The electric oven is equipped with left-tilting means 10 and right-tilting means 11, in the form of a vertical jack resting on the floor, for example.

[0039] The two electrodes form two electric arcs 14. The arcs are free (outside the material, in the air) and there is a transmission of heat by radiation, thanks to the presence of a jet of gas at very high temperature which radiates towards the material to be melted.

[0040] The loading of the material to be treated and the reagents is carried out between the two electrodes 3 and 4. The two arcs 14 are further positioned towards the metal and slag casting orifices, respectively 5 and 6.

[0041] The furnace is a direct current furnace and the two electrodes 3 and 4 are graphite electrodes.

[0042] When manufacturing a carburized ferroalloy from the waste materials mentioned above, the melting temperature of the carburized ferroalloy is in the order of 1250-1350°C, and it is therefore reasonable to heat only to just above this temperature, to save the energy used.

[0043] Furthermore, due to the high proportion, in the treated material, of slag-generating mineral matter, or of elements such as manganese leaving a poorly fusible oxide in the slag, the resulting slag can accumulate and form accretions.

[0044] At temperatures close to the melting temperature of the slag, the latter becomes pasty and tends to form accretions.

[0045] The melting process follows a cycle starting with a metal bath base having a composition close to that of the carburized ferroalloy which is intended to be produced - this bath foot being brought to 1450-1500°C.

[0046] Starting from this bath foot, which represents from a quarter to a third of the metal capacity of the furnace tank, the electrical power is applied and the waste to be treated is continuously charged - in bulk with a particle size of 2 to 150 mm or in the form of agglomerates - pellets, briquettes or extrudates. Melting takes place and the material accumulates in the furnace, being divided between metal and slag.

[0047] This continuous melting phase is stopped when the slag reaches a limit height beyond which it is no longer heated efficiently by the arcs, and the slag is cleaned by tipping it towards the casting door 5. The slag 12 is thus emptied out of the furnace. The molten metal is, however, left in place, and the melting phase is resumed.

[0048] This sequence of continuous waste reduction-melting, interspersed with slag de-slagging(s), is repeated until the height of the metal itself reaches a limit which is the metal capacity of the furnace tank. Then the slag is poured through the slag pouring door 5 and then this time also the metal through the metal pouring channel 6. The pouring of the metal is limited to two thirds of the metal capacity of the furnace, the remaining third being kept in the furnace and constituting the bath base allowing a melting cycle to be restarted under good conditions.

[0049] The process is then repeated as above, by continuously adding material to be treated.

[0050] The central loading of the materials is carried out in the vault of the furnace, so that the materials fall in the middle of the axis connecting the two electrodes. The material loading orifice is located so that the materials fall into the slag between the two arcs.

[0051] The two arcs 14, very stable due to the current supply in direct current mode, are naturally oriented towards the two casting orifices 5 and 6, and repel each other by virtue of the Lenz-Faraday electromagnetic law.

[0052] The stirring elements of the metal-slag bath, consisting of tuyeres or porous bricks placed at the bottom of the hearth, may be one, two or more in number. Their number is to be defined according to the size of the furnace: only one for a small furnace, up to five for a furnace of a few dozen tons. The entire metal-slag exchange surface must be well stirred, to ensure the reduction-fusion reactions and contribute to the homogeneity of the slag.

[0053] [Fig.2] The present invention is well illustrated by a more detailed description of the treatment of raw used industrial catalysts, i.e. not previously roasted, which would have had the effect of reducing their hydrocarbon and sulfur contents. In this case, but also in other cases, the production and management of slag are particularly critical due to the large quantity of slag produced.

[0054] The process begins with a furnace filled to one third of its capacity of molten metal in the tank, with a ferroalloy of nickel and molybdenum, and possibly cobalt FeNiMo(Co), constituting a starting reagent RI qualified as a metal bath base.

[0055] This is heated by the electric oven, and is in liquid form.

[0056] A reagent R2 is then gradually introduced by continuous loading into the furnace tank. This is the material to be treated, which in the case mentioned here consists of used industrial catalysts, mixed upon introduction with lime and iron oxide. The details of the composition of reagent R2 are indicated below.

[0057] The raw used catalysts consist of alumina A12O3 at a level of approximately 60%, in which are trapped approximately 2% by mass of nickel and 8% by mass of molybdenum in the form of sulfides, at a level of almost 10% of sulfur by mass, as well as hydrocarbon pollutants, at a level of approximately 10% of carbon by mass, and approximately 5% of hydrocarbons by mass.

[0058] The catalysts are in the form of porous alumina rods in which Nickel and Molybdenum are trapped. Alumina A12O3 is a refractory oxide, which only melts at high temperatures (over 2000°C), but by adding an equivalent quantity of lime (CaO), the 50 / 50 mixture melts around 1450°C. Both alumina and lime are refractory oxides (i.e. with a high melting temperature) and non-reducible, i.e. it is very difficult to dissociate the aluminum from the oxygen in the oxide A12O3 - unlike, for example, iron oxide Fe2O3 which is easily reduced to metallic iron, with carbon, by the reaction Fe2O3 + 3 C -> 2 Fe (metal) + 3 CO (gas).

[0059] These raw used catalysts are melted in the tank with two main additions included in the reagent R2, namely approximately 300 kg of iron oxide per tonne of catalyst to be treated, to recover the nickel and molybdenum in a carburized ferroalloy FeNiMoC, and approximately 600 kg of lime CaO to form a CaO-A12O3 slag fusible at 1500°C, allowing the extraction of the majority of the sulphur present in the material to be treated. The total of the components thus indicated and forming the reagent R2 is 190%, relative to the quantity of catalysts to be treated.

[0060] As shown in [Fig.2], the raw spent catalysts, iron oxide and lime form the reactant R2.

[0061] The actual melting phase, denoted E1, is then carried out, with continuous loading of the catalysts to be treated and the indicated treatment reagents, namely iron oxide and lime.

[0062] This phase is almost continuous, but it must be interrupted each time the slag height reaches a critical level, from which the slag is no longer homogeneous, and begins to solidify either at the wall of the furnace, or opposite the hole or the door of cleaning dedicated to the removal of slag from the furnace. Most of this slag is then poured, constituting a PI product, then the loading of the RI reagent and the El melting continue.

[0063] Molybdenum-rich dust is released and constitutes a product P3. It is recycled during a recycling process E3, for example by extraction in a post-combustion circuit (combustion of impurities) of the fumes and filtration of these fumes. The molybdenum-rich dust P3 is reintroduced at the end of the recycling process E3 into the reagent R2.

[0064] When the metal reaches the metal capacity of the furnace tank, a refining E2 is carried out by adding reagents R3 which are desulfurization and dephosphorization agents, to adjust the sulfur and phosphorus contents of the ferroalloy, which constitute impurities; then the metal alloy formed is poured, which constitutes a product P2, while retaining a bath base of one third of the tank capacity, which constitutes a product P4, and the next phase can begin, the product P4 constituting the reagent RI indicated previously.

[0065] The overall balance of the fusion is written 1 ton of raw used catalysts plus 300 kg of iron oxide plus 600 kg of lime, giving 280 kg of ferroalloy plus 1.3 tons of slag. The ferroalloy includes molybdenum and nickel, and possibly cobalt.

[0066] We therefore obtain 4 to 5 times more slag by mass than metal, and as the densities are 7 for the metal and approximately 2 for the slag, we form in volume nearly 15 times more slag than metal.

[0067] To clarify and quantify the importance of controlling the temperature of the slag, we take the example of a 20-tonne furnace, whose tank has an internal diameter of 2.5 meters, i.e. a bath surface area of ​​approximately 5 m2.

[0068] This furnace is loaded at a ratio of 6 tonnes per hour of catalysts, as well as excess Fe oxide and lime. 1.68 tonnes of metal and 7.8 tonnes of slag are formed per hour of melting, or 5 cm of metal and approximately 78 cm of slag.

[0069] If the slag temperature is controlled, especially at the slag taphole, it is possible to work and heat with stable electric arcs 14 up to 40 cm of slag height. This implies that the slag must be poured every 30 minutes to clean the furnace - and in this controlled case, a pour takes about 5 minutes.

[0070] In an uncontrolled case, i.e. with slag accretions on the walls and / or the tap hole, the metal and slag assembly must be overheated, i.e. a mass of 10 to 22 tonnes to be overheated by 50°C, i.e. an energy expenditure of 150 to 300 kWh for each overheating, and an additional loss of production time of 5 to 10 min, including observation and restart times.

[0071] Over a complete cycle, the maximum metal height is reached, i.e. 20 tonnes of metal after 7.5 hours, or 12.6 tonnes of metal produced in addition to the 7 tonnes of bath foot, after 15 slag pours - one every Uz hour. The contribution of maintaining a moderate slag temperature and limited overheating at the taphole - which the proposed furnace configuration allows - is to avoid losing 15 times 5 to 10 min or 75 to 150 minutes, and to consume 15 times 300 kWh or 4.5 MWh of electrical energy per sequence or melting cycle.

[0072] A melting cycle comprises 7.5 hours of melting and 75 minutes of slag pouring, 115 minutes for refining and pouring the metal, i.e. a total of 10 hours. The loss of 5 to 10 minutes for each pour therefore represents an extension of the cycle from 75 to 150 minutes to 600 minutes, i.e. 12.5 to 25% loss of production time.

[0073] This industrial advantage is found in the case of the melting of alkaline (zinc-manganese dioxide), saline (or dry batteries, or Leclanché batteries) and zinc-air (rechargeable or non-rechargeable) batteries, accumulators or other waste whose melting produces a lot of slag. These batteries contain a lot of iron and zinc, both reducible, and also manganese oxide in the form of MnO2 or MnO. However, this manganese oxide is refractory (it only melts at high temperatures) and is not completely reducible. No iron oxide is added, but lime, limestone or dolomite are added to make the slag fusible. A significant portion, around 20%, of the manganese remains in the form of MnO oxide in the slag. We know how to resell this slag rich in Mn, in particular because it contains lime, to companies using manganese ore.The addition of lime thus not only helps to make the slag fusible, but also to make it recoverable due to the presence of lime in the manganese oxide. We therefore recover the manganese in the slag, and the iron and zinc and part of the manganese in the ferroalloy.

[0074] [Fig.3] [Fig.3] is a top view of the furnace. Electrodes 3 and 4 are visible, as well as the slag pouring door 5 and the metal pouring channel 6. The slag 12 is visible above the metal alloy, which occupies the bottom of the tank. A flue 16 is present in the vault, and opens above the slag through an orifice 16a. The tank has an internal diameter of 2.5 meters, and the distance between the slag pouring door 5 and the electrode closest to it, namely electrode 4, is of the order of 70 cm.

[0075] The figure shows the vertical plane P in which both the ends 3a and 4a of the electrodes and the slag pouring gate 5 are located, as well as, advantageously, the contact point 2a in the slag of the solid material to be treated, which falls from the loading opening 2. The metal flow channel 6 is also located in the plane P, but this is not as essential. Thanks to the focusing of the energy in this plane, the slag is particularly fluid near the slag pouring gate 5, which allows cleaning to be carried out without expense excessive energy, nor waste of time.

[0076] The ends of the electrodes 3a and 4a are not necessarily at the same height above the slag surface (i.e. the same altitude), but it is advantageous if they are, because the construction of the device and the control of the process are then simpler.

[0077] By closely associating the slag pouring orifice and one of the direct current electrodes, the material can be heated intensely specifically in the pouring zone. The rest of the slag may remain pasty, but the area close to the orifice is liquid.

[0078] The preferred embodiment for slag extraction is the tilting of the furnace tank with the slag passing through an overflow at the edge of the tank. Alternatively to this tilting and overflow system, an outlet through an overflow without tilting can be used. As a final alternative, a submerged door (also called a taphole) can also be used, which is opened and closed successively or even periodically.

Claims

Claims

1. A method for recycling metal comprising melting (El) in an electric electrode furnace tank (3, 4) a material to be treated (R2) comprising metal residues to which lime or a source of lime is added, pouring through a liquid discharge path (5) to the outside of the tank a slag (12, PI) formed during the melting (El) and recovering from the tank a ferrometallic alloy produced (13, P2) also during the melting (El), the method comprising supplying direct current to two electrodes (3, 4) of the furnace placed above the molten material (12, 13), an opening in the tank of the liquid discharge path (5) for said pouring and the ends (3a, 4a) of the two electrodes (3, 4) for the direct current supply being placed in a common vertical plane (P),and a drop point (2a) in the molten material of the material to be treated is placed in the common plane (P) of the opening in the tank of the liquid discharge path (5) and of the ends (3a, 4a) of the two electrodes (3, 4), characterized in that it comprises a metal flow channel (6) placed in the common plane (P), the channel being located opposite the liquid discharge path (5), and in addition a stirring of the molten material (12, 13) by injection of gas through the bottom wall of the tank.,

2. Method for recycling metal according to claim 1, characterized in that it comprises a progressive loading of the material to be treated (R2) into the tank and several successive evacuations of the slag (PI), separated by melting phases, before an evacuation of the ferro-metallic alloy produced (P2).

3. Method for recycling metal according to claim 1 or 2, characterized in that the pouring of slag (PI) and the recovery from the alloy tank (P2) are carried out by tilting the tank respectively on a first side and a second side opposite the first side.

4. Method for recycling metal according to one of claims 1 to 3, characterized in that after the recovery from the tank of the produced ferrometallic alloy (P2), a bath base of produced ferrometallic alloy (P4) is kept in the tank and a material to be treated (R2) comprising metallic residues is again introduced to continue the treatment.

5. A method of recycling metal according to one of claims 1 to 4, ca- characterized in that the material to be treated (R2) comprises at least nickel, molybdenum, iron or zinc.

6. Method for recycling metal according to one of claims 1 to 5, characterized in that lime, limestone or dolomite are added to the material to be treated (R2) in order to obtain in the molten material (12, 13) a ratio of mass of lime plus mass of magnesia to mass of alumina plus mass of silica strictly greater than 1.

2.

7. Method for recycling metal according to one of claims 1 to 6, characterized in that the material to be treated (R2) consists of raw used catalysts.

8. Method for recycling metal according to one of claims 1 to 6, characterized in that the material to be treated (R2) consists of used batteries or accumulators.

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