US3677926A - Cell for electrolytic refining of metals - Google Patents

Abstract

A FUSED SALT CELL FOR THE ELECTROLYTIC REFINING OF METALS, HAVING A LOWER MOLTEN METAL CATHODE, AN UPPER MOLTEN METAL ANODE, A FLUX-IMPREGNATED POROUS DIAPHRAGM DISPOSED BENEATH AND IN CONTACT WITH THE ANODE AND A LAYER OF MOLTEN FLUX ON TOP OF THE ANODE.

Description

y 18, 1972 w. D. DAVIS CELL FOR ELECTROLYTIC REFINING OF METALS Filed June 16, 1970 2 Sheets-Sheet 1 July 18, 1972 w. D. DAVIS 3,677,926

CELL FOR ELECTROLYTIC REFINING OF METALS Filed June 16, 1970 2 Sheets-Sheet 2 VI -s,:-- 29 United States Patent Filed June 16, 1970, Ser. No. 46,659 Int. Cl. C2211 3/02, 3/20 US. Cl. 204-243 R 6 Claims ABSTRACT OF THE DISCLOSURE A fused salt cell for the electrolytic refining of metals, having a lower molten metal cathode, an upper molten metal anode, a flux-impregnated porous diaphragm disposed beneath and in contact with the anode and a layer of molten flux on top of the anode.

This invention is concerned with the electrolytic refining of metals by the fused salt method, i.e. by the method which involves the use of a fused salt electrolyte (hereinafter termed the flux) in conjunction with liquid metal electrodes. In carrying out this method pure metal is used as the cathode and impure metal as the anode.

This method has considerable advantages over electrolytic refining with the use of solid electrodes and an aqueous electrolyte. Firstly much higher current densities can be used and consequently a considerably smaller plant is required. Secondly, certain difiiculties which arise with solid electrodes and an aqueous electrolyte are avoided, namely the formation of dendritic cathode deposits and anode slime. Thirdly, the necessity for casting special anodes and cathode starting sheets is avoided. Fourthly, the liquid metal formed at the cathode during electrolysis may be tapped Without drag-out of electrolyte. Nevertheless a problem arises in providing for effective separation of the liquid constituents of the refining cell, namely the anode, the cathode and the intervening flux.

In the case of aluminium it is possible to use a cell containing a bottom layer of a liquid heavy aluminium alloy as the anode, a superposed layer of molten flux and a top layer of pure aluminium as the cathode. With many metals, however, it is impossible to formulate a flux of density intermediate between that of the anode and the cathode so that this form of flotation cell cannot be used.

Another possibility is to utilize as the cell two concentric circular vessels, the inner one containing the liquid metal anode and the outer one of liquid metal cathode, and to provide a layer of flux floating on both electrodes. In this case the current passes through the flux disposed above the wall separating the electrodes. A cell of this construction cannot, however, be made of sufficient size for effective commercial operation.

A third possibility is to utilize between the flux and the electrodes porous rigid diaphragms of ceramic material which are penetrated by the flux but not by the metal of the electrodes. Owing, however, to the high temperatures at which the cell must be operated, 500 C. or more, such rigid diaphragms are liable to rapid failure from cracking which is probably caused by thermal expansion. Also considerable problems arise in providing effective seals at the edges of the diaphrag rns.

The invention provides a fused salt method for the electrolytic refining of metals, in which the refining operation is carried out in a cell containing a lower molten metal cathode, an upper molten metal anode, a flux-impregnated porous diaphragm disposed beneath and in ice contact with the anode and a layer of molten flux on top of the anode.

In early experiments on the electrolytic refining of lead the diaphragm rested on a block of graphite which formed the cathode, and the liquid metal deposited electrolytically on it was removed from the cell via a system of grooves and channels in the graphite. Under these conditions, the diaphragm was rapidly penetrated by molten metal deposited electrolytically in its pores, and this penetration continued until the graphite cathode and the molten metal anode were short-circuited by a continuous metallic path. At a later stage a block of graphite formed with a well was used to contain the molten metal cathode, the metal being drawn off from the cathode by a rotatably mounted siphon tube, by rotation of which the level of the pool of cathode metal could be adjusted. The diaphragm was supported above the well by perforated ceramic plates and around the well by the raised, outer, portion of the graphite block, but insulated from it by thin ceramic tiles. Difiiculties were experienced sometimes from penetration due to flaws in this insulation, but such cells did function when the cell was operated under conditions such that a shallow layer of molten flux was maintained between the cathode and the undersurface of the diaphragm. This precaution was continued after a modified cell had been developed in which the well containing the cathode was formed in a body of non-electrically-conductive ceramic material instead of in a block of graphite.

Recent further experiment has led to the surprising discovery that, when the well containing the cathode is formed in a block of non-conducting material it is possible to operate the cell with the siphon tube adjusted to a position corresponding to contact between the cathode metal and the undersurface of the diaphragm, and therefore under conditions of minimum internal resistance and power consumption, without any tendency to deposition in the diaphragm of metal travelling towards the cathode.

This discovery arose from an experiment in which the internal resistance of the cell was measured at successive intervals of time with the siphon tube blocked but so adjusted that the level of the cathode metal was below the undersurface of the diaphragm.

Under these conditions, the resistance fell progressively as the level of the cathode rose until it reached a minimum value at which it remained during continued running of the cell. This minimum value of the resistance corresponded to contact or substantial contact of the cathode metal with the diaphragm, metal which was added to the cathode as the result of continued electrolysis thereafter moving upwardly around the peripheral portion of the diaphragm. Electrolysis continued satisfactorily Without any deposition of metal within the diaphragm until arcing ultimately occurred due to cathode metal running back into the anode.

In another experiment, with the cell run with the siphon tube open and in a position such that the cathode metal should be in contact with the undersurface of the diaphragm, it was found that the level of metal in the siphon tube dropped slightly when the current was switched off and that the level of the molten metal in the siphon tube once more rose to the brim immediately the current was switched on again. This suggests that in operation the level of the cathode pool is very slightly lower than it should be for hydrostatic equilibrium. Possibly electrical or other forces may be effective while the cell is running to maintain a very slight spacing between the cathode metal and the undersurface of the diaphragm. Such spacing must, if it exists, be microscopic having regard to the fact that the surface area of the cathode metal pool is vastly greater than the surface area of the column of metal in the siphon tube and the rise and fall of the metal in the siphon tube when the cur rent is switched on and oif is only of the order of inch. An electrical action of some kind probably aifords the explanation having regard to the high current density 2000-3000 amps. per sq. ft., which is needed to keep the flux and the anode metal molten.

Preferably the diaphragm is flexible and consists of a felt of ceramic fibres. Such felts are commercially available, being supplied by Morganite Refractories (Sales) Ltd. under the trademark Triton and have hitherto been used for thermal insulation. Such material has a density of 6-8 lbs/cu. ft. and the pore size is small enough to hold a head of several inches of molten lead. The Triton felt which has been successfully used contains alumina and silica in the approximate molecular ratio of 1:2.

The method according to the invention is not only applicable to the refining of metals in the true sense, in which a metal containing minor amounts of impurities is used as the anode, substantially pure metal being recovered at the cathode and the anode becoming progressively enriched with impurities but also to what might more correctly be termed partition of metals, in which the anode is an alloy of two metals containing a substantial amount of both metals and both metals can be recovered in a substantially pure state, one at the cathode and the other at the anode.

The invention is particularly concerned with the refining of lead and lead alloys, e.g. alloys of lead and antimony, bismuth or tin. It can be used by primary producers in the production of pure lead (e.g. by removal of bismuth from the lead) but it is also useful in the working up of secondary products, e.g. of lead-antimony alloys from scrap electrical batteries. In such a case, it is possible for example to produce almost pure antimony by depleting the anode of lead to an extent such that its lead content is 2% or less, While at the same time obtaining lead of moderate purity at the cathode. By utilising as the anode a printing metal alloy of lead, antimony and tin it is possible to recover at the cathode a lead-tin alloy which is almost entirely free from antimony. Other examples are the refining of tin by removing impurities such as arsenic, iron, copper, antimony and bismuth, the refining of zinc by separating impurities such as lead, cadmium, tin and copper, the separation of zinc and cadmium from a zinc-cadmium alloy used as the anode and the separation of gold from scrap solder.

In the case when the anode is of lead or an alloy of lead a suitable flux consists of 79% lead chloride, potassium chloride and 6% sodium chloride, proportions being by Weight. If desired lithium chloride may be added to the flux to improve its electrical conductivity.

It has been found that a diaphragm of the above-described felt material tends to develop weakness which leads ultimately to failure and that this arises from chemical reaction between the material of the diaphragm and the oxide ions in the flux which are derived, in the case of lead, from lead oxide introduced as dross on the metal used for topping up the anode. Such oxide ions can be eliminated by incorporating in the flux a salt, usually in an amount of 2-5% by weight, of a metal which forms an oxide which is insoluble in the flux. One suitable pre cipitating salt is stannous chloride.

When the material to be refined is a lead alloy obtained by smelting secondary lead, the alloy often contains sufficient tin to result in the formation in the flux, by reaction with the lead chloride, of sufficient stannous chloride to produce the required precipitating action. In such a case no separate addition of stannous chloride to the flux is necessary, because part of the tin in the flux will be deposited with the lead at the cathode and the supply will be replenished by fresh tin added as part of the impure lead added to top up the anode. A sufiicient concentration of stannous chloride in the flux will thus be maintained.

Any tin the flux will transfer with the lead to the cathode but this will be immaterial when the prime object is .to recover antimony at the anode. When, however, the prime object is to obtain pure lead at the cathode it is necessary to use for precipitation a salt of a metal which will not deposit from the flux on the cathode. A suitable salt for such a case is cerium chloride. j

Other metals which can be used to precipitate oxide are chromium, uranium, thorium and magnesium. Indeed the first three of these are believed to be more effective than tin or cerium.

Cerium is most conveniently introduced into the flux by addition to it of cerium chloride. The other metals may most conveniently be added as metal to the molten flux, when they react to precipitate lead, e.g. as follows:

A cell suitable for use in carrying out the invention is illustrated in the accompanying diagrammatic drawings, in which:

FIG. 1 is a vertical section through the cell,

'FIG. 2 is an enlarged vertical section showing the cathode conductor and draw-off tube, and

FIG. 3 is a section on the line 1II-HI in FIG. "I.

The cell shown in the drawing is a square cell measuring 13" x 13" and has an outer mild steel sheet metal casing 10, of height 7 /2", surrounded by a water cooling coil 11. Such a cell is operated at a current of 800-850 amperes and a voltage of 4-5 volts. Within the casing 10 are four blocks 12 of ceramic material, which are shaped to form a well 13 containing the molten metal cathode 14 and a superposed layer 15 of molten flux. The layer 15 of flux is in contact with the undersurface of a diaphragm 16 of the above-mentioned ceramic felt Which is impregnated with the flux. The main portion of the diaphragm 16 extends horizontally, being supported on spaced bars 17 of ceramic material and its edges are dished and supported on correspondingly-shapedformations 18 on the blocks 12. To permit of escape of gases generated from the flux upon electrolysis "two opposite edges of the diaphragm are maintained spaced from the blocks 12 by ribs 9 (FIG. 3) on the blocks. The molten anode metal 19 is contained within the dish formed by the diaphragm 16 and is covered by a layer 20 of molten flux. The layer 20 of flux is in contact with the diaphragm 16 to ensure that the diaphragm will remain saturated with flux and serves to protect the anode metal 19 from atmospheric oxygen. The cell has a removable lid 21 of asbestos cement composition, having an inlet opening 22 for topping up the anode and carrying a graphite anode conductor 23, a sheathed thermocouple 24'and a graphite baffle plate 26, supported by rods 25, for protecting the diaphragm 16 from the direct impact of added anode metal. Anode metal is withdrawn from the cell, when required, by suction.

The cathode conductor is constituted by a stainless steel tube 27, to which is welded a disc 28 carrying a peripheral water cooling coil 29 which is connected to the DC supply. A packing 30 of asbestos cement composition is interposed between the disc 28 and the Wall 10. Packing 31 of ceramic fibre is provided around the outer end of the tube 27 and also between the wall 10 and the blocks 12. A siphon tube 32, rotatably mounted in the tube 27, provides for continuous withdrawal of cathode metal from the cell. By rotation of the tube 32 the level of the pool of cathode metal 14 in the cell can be adjusted. In use a flame may be played on to the tube 32 to counteract heat loss by radiation. The heat generated in the tube 27 by passage of current through it ensures that the cathode metal within it will remain molten. Leakage of flux and cathode metal from the cell is prevented by solidification of the flux near the water cooled wall 10 and of cathode metal near the Water cooled flange 29.

In the following examples a cell as just described was used with the exception of Examples 4 and in which a smaller laboratory scale apparatus operating on the same principle was utilised. In these examples percentages are by weight.

A cell, containing a flux consisting of 73% 'PbCl 16% KCl, 6.5% NaCl and 4.5% MgCl obtained by reacting 1 part of magnesium metal with 100 parts of the ternary flux was started with a pure lead anode and a pure lead cathode. After six hours, during which the anode was fed with pure lead and during which lead was produced at the cathode at a rate of 3.1 kgs. per hour, the feed was replaced by a lead-antimony alloy containing of antimony. When a total of 5.86 kgs. of antimony had been added to the anode pool, the cathode metal contained 0.0012% antimony and the anode pool contained 59.5% antimony. Electrolysis was continued, adding no further anode metal, until the anode contained 99% of antimony, this point being indicated by a sharp fall in the current through the cell. The cathode metal at this point contained 0.035% of antimony. The experiment was continued by adding lead to the anode pool to restore the anode pool to its original level and electrolysis was resumed. The anode level was maintained for some time by adding lead at the rate at which lead was produced at the cathode. Feeding the anode was then stopped for a second time and electrolysis continued until the anode contained 99% of antimony. The cathode metal then contained 0.012% antimony.

A cell containing a flux consisting of 74% PbCl 14% KCl, 5 /2% NaCl, .and 6 /2% CeCl was started with a pure lead cathode and an anode consisting of a leadantimony alloy containing 27% antimony. After 8.0 kgs. of antimony had been gradually introduced by addition of this alloy, addition of fresh anode metal was halted and electrolysis continued. When the anode contained 98.7% antimony, the cathode lead was found to contain 0.045% of antimony. Lead was then added to the anode pool and a further 1.3 kgs. of antimony was added in the form of a lead-antimony alloy containing 10% of antimony. Addition of fresh anode metal then ceased and electrolysis was continued until the anode metal consisted of substantially pure antimony. At this point, 4.2 kgs. of anode metal was removed by suction. This contained 98.6% of antimony and the cathode metal was found to contain 0.009% of antimony.

A cell was started with a flux consisting of 79% PbCl KCl and 6% NaCl which had been allowed to react with one-tenth of its weight of tin, and with a cathode and an anode both consisting of a lead-tin alloy containing 3% tin. Electrolysis was continued and the anode was replenished with a printing metal alloy containing 11.3% antimony, 3.3% tin, the balance lead. The addition of anode alloy was halted when 8.8 kgs. of antimony had been introduced in the form of this alloy. Electrolysis was then continued until the anode contained 99% of antimony. At this point the cathode contained 0.007% of antimony and 3.6% of tin. Part of the anode metal (3.1 kgs.) was removed by suction and 6.8 kgs. of a lead-tin alloy containing 3.6% tin was added. Electrolysis was continued replenishing the anode with an alloy containing 10.9% antimony, 3% tin, the balance lead until the cell contained 7.9 kgs. of antimony. Electrolysis was then continued with no addition of anode metal until the anode contained 99% of antimony and the cathode 0.003% antimony. This example illustrates the recovery from a printing metal alloy of a lead-tin alloy.

In an experiment on the separation of zinc and cadmium, 150 gms. of a zinc-cadmium alloy containing 6 16.9% cadmium was electrolysed using a flux consisting of 42% ZnCl- 50% KCl, and 8% NaCl. After electrolysis the anode metal weighed 58 gms. and consisted of 41.0% cadmium balance zinc and the cathode metal weighed 77 gms. and consisted of zinc containing 1.1% of cadmium.

Original anode, Cathode, percent; percent;

A cell as described above was started using commercial lead both in the cathode well and for the anode feed. The flux consisted of the ternary eutectic of lead chloride (79% by Weight), potassium chloride (15%) and sodium chloride (6%) which had been reacted with metallic chromium to introduce some chromium chloride to precipitate the oxide ion, and had the approximate composition 71% lead chloride, 16% potassium chloride, 6 /2% sodium chloride, and 6 /z% chromium chloride. After a period of electrolysing lead, the anode feed was changed to a scrap solder alloy containing 0.32% of gold together with some silver and 60.9% of tin. There was, as expected, an exchange reaction whereby some tin chloride entered the flux and a corresponding amount of lead entered the molten metal phases. A total of 495 lbs. of this solder was fed to the anodes and there was produced 541 lbs. of cathode solder containing less than 0.001% of precious metals and 48.8% of tin; and a final anode residue weighing 5.7 lbs. and containing 25.7% of silver (27.1% precious metals).

What I claim as my invention and desire to secure by Letters Patent is:

1. An electrolytic cell for use in the electrolytic refining of metals by the use of liquid metal electrodes and a fused salt electrolyte, said cell comprising a body of solid, electrically non-conducting material defining a well, a pool of cathode metal in said Well, a porous diaphragm of flexible ceramic material adapted to be impregnated with electrolyte and disposed on said body, said diaphragm having a central portion extending substantially over said well to cover the same and dished edge portions supported on said body, a pool of anode metal resting on and supported by the upper surface of said diaphragm, an anode conductor in contact with the anode metal, a cathode conductor in contact with the cathode metal, and an upturned siphon tube for removing cathode metal from the cell, said siphon tube being angularly adjustable to vary the level of cathode metal in the well.

2. A cell as claimed in claim 1, which includes spaced supports beneath the central portion of the diaphragm.

3. A cell as claimed in claim 1, wherein the diaphragm is constituted by a felt of ceramic fibres.

4. A cell as claimed in claim 1, wherein the cathode conductor is a tube within which the siphon tube is rotatably mounted.

5. A cell as claimed in claim 4, in which said tube carries a disc external to the cell and surrounded by a peripheral cooling coil.

gold and 1.4% of externally cooled casing surrounding said body.

References Cited UNITED STATES PATENTS Hoopes 204-243 R Auerbach 204--66 X Ingeberg 204-66 Dittmer 204243 R 8 3,383,294 5/1968 Wood 204-245 X 3,578,580 5/1971 Schmidt-Hatting et a1.

1 204243 GERALD L. KAPLAN, Primary Examiner D. R. VALENTINE, Assistant Examiner US. Cl. X.R. 204-245, 250,251

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