Patent Application
20100181205
The present invention relates to a method for treating waste containing precious metals as well as a device for implementing this method.
The increase in the use of calculating machines, mobile phones, electronic apparatuses and other short life high-tech apparatuses generates an increasing amount of waste containing rare and precious metals. This situation poses the problem of recovering and treating the metals contained in this waste. Such waste is thus a real source of non-ferrous, rare and precious metals.
Electronic waste is presently collected, exported and treated in large industrial complexes for non-ferrous metals often requiring several plants in succession for extracting lead, copper and zinc where these high value materials are diluted in flows of raw materials of mining or secondary origin. Present methods are therefore optimized in order to produce lead, copper or pure zinc in large amounts, but they are poorly adapted for producing rare and precious metals present in small amounts. It is therefore desirable to design a method for treating waste with which a large proportion of the precious metals contained in the waste may be recovered.
A first object of the invention is a method for treating waste containing precious metals, comprising the following successive steps:
contacting the waste with a molten lead-based composition;
skimming the mixture obtained; and
refining the skimmed mixture by electrolysis so as to recover the precious metals.
According to a particular embodiment, in the skimming step, residues are recovered, which are treated by:
contacting the residues with a second molten lead-based composition;
skimming the mixture obtained; and
recovering the skimmed mixture in order to provide at least one portion of the portion of the aforementioned molten lead-based composition.
According to a particular embodiment, the step of refining the skimmed mixture comprises the following sub-steps:
casting the skimmed mixture into anodes;
electrolysis of a fluorosilicic acid solution by using said anodes; and
recovering anodic sludges containing the precious metals.
According to a particular embodiment, the aforementioned method comprises before, simultaneously or after the step of recovering anodic sludges, the step of:
recovering lead and possibly tin cathode deposits of in order to provide at least one portion of the molten lead-based composition and/or of the second molten lead-based composition.
According to a particular embodiment, the step of refining the skimmed mixture comprises, after the sub-step of recovering anodic sludges, the following additional sub-steps:
melting the recovered anodic sludges in the presence of oxygen,
skimming the molten anodic sludges; and
casting the molten and skimmed anodic sludges into ingots.
According to a particular embodiment, each molten lead-based composition comprises 0-50% of tin, preferably 0-20% of tin.
According to a particular embodiment, the aforementioned method comprises, prior to the step of contacting the waste with a molten lead-based composition, the step of:
extracting copper from the waste by selective dissolution.
According to a particular embodiment, the copper extraction step comprises the following sub-steps:
selectively dissolving the waste in the presence of sulfuric acid, iron sulfate and oxygen;
treating the obtained solution by filtration and/or electrolysis and/or precipitation;
recovering copper on the one hand and other metal impurities on the other hand.
According to a particular embodiment, the aforementioned method comprises, prior to the copper extraction, the following step:
combustion of the waste by pyrolysis, producing carbonaceous gases; and optionally
post-combustion of the carbonaceous gases.
According to a particular embodiment, the aforementioned method further comprises a preliminary step of milling the waste and/or analyzing milled waste. According to a particular embodiment, the precious metals comprise one or more metals selected from gold, silver, platinum, palladium, rhodium, ruthenium, iridium, osmium and mixtures thereof. According to a particular embodiment, the waste is selected from catalytic exhaust mufflers and electronic waste such as electronic cards. According to a particular embodiment, more than 90% by mass, preferably more than 99% by mass of the precious metals contained in the waste are recovered according to the aforementioned method. According to a particular embodiment, the supernatant materials comprise ceramics, glass fibers and/or ferrites.
Another object of the invention is an installation for treating waste containing precious metals, comprising:
at least one container for covering with lead;
a molten lead-based composition feed line, connected to the inlet of the lead covering container;
a pretreated materials feed line, connected to the inlet of the lead covering container;
skimming means associated with the lead covering container;
a skimmed mixture withdrawal line, connected to the outlet of the lead covering container;
means for refining the skimmed mixture by electrolysis, fed by the skimmed mixture withdrawal line; and
a precious metals withdrawal line, connected to the outlet of the means for refining the skimmed mixture.
According to a particular embodiment, the aforementioned treatment installation further comprises:
a skimming residues withdrawal line, connected to the outlet of the skimming means;
at least one additional lead covering container, fed by skimming residues withdrawal line on the one hand, and by an additional molten lead-based composition feed line on the other hand;
additional skimming means associated with the additional lead covering container; and
an additional skimmed mixture withdrawal line, connected to the outlet of the additional lead covering container and feeding the molten lead-based composition feed line.
According to a particular embodiment, the means for refining the skimmed mixture by electrolysis comprise:
means for casting anodes;
means for Betts electrolysis; and
means for recovering anodic sludges.
According to a particular embodiment, the means for refining the skimmed mixture by electrolysis comprise:
means for recovering lead-tin, feeding the additional molten lead-based composition feed line.
According to a particular embodiment, the aforementioned treatment installation further comprises:
copper extraction means, one outlet of which is connected to the pretreated materials feed line; and
a primary materials feed line, connected to the inlet of the copper extraction means.
According to a particular embodiment, the copper extraction means comprise:
at least one selective dissolution container fed by the raw materials feed line;
a depleted electrolyte feed line, connected to the inlet of the selective dissolution container;
electrolysis means,
means for transferring rich electrolyte, connecting the selective dissolution container to the electrolysis means;
means for stripping the cathodes; and
means for recycling depleted electrolyte, connected to the outlet of the electrolysis means.
According to a particular embodiment, the aforementioned treatment installation further comprises:
means for pyrolysis of the waste, connected to the outlet of the raw materials feed line;
a waste feed line feeding the pyrolysis means; and optionally
a gas exhaust line at the outlet of the pyrolysis means and feeding post-combustion means.
According to a particular embodiment, the aforementioned treatment installation further comprises:
milling and analyzing means fed by a raw waste feed line and feeding the waste feed line.
According to a particular embodiment, the aforementioned method is applied in the aforementioned installation. According to a particular embodiment, the aforementioned installation is intended for implementing the aforementioned method.
The present invention makes it is possible to overcome the drawbacks of the state of the art. It more particularly provides a specific method for treating and recovering waste containing precious metals, which harmoniously combines metallurgical sequences and avoids dilution of the contained metals in a production flow of primary metals. The invention also provides a single installation for separating the constituents from the waste and in particular for recovering precious metals.
According to certain particular embodiments, the invention also has the advantageous features listed below.
The method according to the invention is very flexible and it may be adapted to foreseeable changes in the composition of electronic cards.
The method according to the invention does not have the drawback of using the customary technique of extracting precious metals by oxidization of lead, the so-called cupellation operation, followed by the operation for reducing lead oxide.
With the invention, it is possible to retain metals in the metal form as much as possible: precious metals are kept in the metal form throughout the method, and the lead and tin are kept in the metal form right up to step (d) included. This makes it possible to minimize the carrying away of the precious metals by metal oxides.
With the invention, the precious metals may be collected at a single outlet.
The invention may be applied with a controlled environmental impact.
The invention is now described in more detail and in a non-limiting way in the following description.
Waste Treatment Installation
Referring to
According to an embodiment, the waste feed line 1 feeds pyrolysis means 2. At the outlet of the pyrolysis means 2, provision is made for means for feeding primary materials 6, which feed copper extraction means 37. At the outlet of the copper extraction means 37, provision is made for a pre-treated materials feed line 14, which feeds a lead covering container 15. This embodiment is particularly well adapted to treating used electronic cards.
According to an alternative, the pyrolysis means 2 are absent, and the waste feed line 1 directly feeds the copper extraction means 37 (in this case it is considered that the waste feed line 1 coincides with the primary materials feed line 6). According to another alternative, the copper extraction means 37 are absent, and the primary materials feed line 6 directly feeds the lead covering container 15 (in this case it is considered that the primary materials feed line 6 coincides with the pre-treated materials feed line 14). According to still another alternative, both the pyrolysis means 2 and the copper extraction means 37 are absent, and the waste feed line 1 directly feeds the lead covering container 15 (in this case it is considered that the waste feed line 1 and the pre-treated materials feed line 14 coincide). This alternative is particularly adapted to the treatment of used catalytic exhaust mufflers, since the latter practically contain no copper.
When they are present, the outlet of the pyrolysis means 2 may be connected to a gas exhaust line 4 which may feed postcombustion means 5. When they are present, the copper extraction means 37 may include a selective dissolution container 7 fed at the inlet by the pre-treated materials feed line 6 and fed by a depleted electrolyte feed line 11 on the other hand. The pre-treated materials feed line 14 is then connected to the outlet of the selective dissolution container 7, while a rich electrolyte transfer line 8 feeds electrolysis means 9. Means for stripping the cathodes 13 are provided in association with electrolysis means 9, and a depleted electrolyte recycling line 10 is provided at the outlet of the electrolysis means 9. This depleted electrolyte recycling line 10 may feed the depleted electrolyte feed line 11 and/or a depleted electrolyte treating line 12.
The lead covering container 15, which is fed by the pre-treated materials feed line 14, is also fed by a molten lead-based composition feed line 24. Skimming means 16 are associated with the lead covering container 15. At the outlet of the lead covering container 15, provision is made for a skimmed mixture withdrawal line 21, which feeds means for refining the skimmed mixture 36. A precious metals withdrawal line 38 is provided at the outlet of the means for refining the skimmed mixture 36.
Provision may also be made for a skimming residues withdrawal line 17 at the outlet of the skimming means 16, which may feed an additional lead covering container 18. This additional lead covering container 18 is then also fed by an additional molten lead-based composition feed line 31.
Additional skimming means 19 are provided in association with the additional lead covering container 18 and an additional skimmed mixture withdrawal line 22 is provided at the outlet of the additional lead covering container 18. This additional skimmed mixture withdrawal line 22 may, just like the skimmed mixture withdrawal line 21, feed the means for refining the skimmed mixture 36. But according to a preferred alternative, the additional skimmed mixture withdrawal line 22 feeds the molten lead-based composition feed line 24. A complementary source of molten lead-based composition 23 may optionally be provided in order to complete this feed. An additional skimming residues withdrawal line 20 may be provided at the outlet of the additional skimming means 19.
The means for refining the skimmed mixture 36 more specifically provide means for casting anodes 25, a system for transferring anodes 26, Betts electrolysis means 27. At the Betts electrolysis means 27, provision is made for means for recovering lead-tin 29 and means for recovering anodic sludges 28. The means for recovering lead-tin 29 feed (optionally together with a fresh lead feed line 30 which may be provided as an option) the additional molten lead-based composition feed line 31.
The means for recovering anodic sludges 28 feed melting means 33, which are further provided with an oxygen feed 32. Final skimming means 34 are provided in association with the melting means 32. The precious metals withdrawal line 38 is connected to the outlet of the melting means 32, which also include a residues discharging line 35.
Now referring to
The first secondary conveyor 106 feeds a coarse mill 109. A fine mill 111 is also provided, fed by the second secondary conveyor 107 on the one hand and by a transfer line 110 on the other hand stemming from the outlet of the coarse mill 109. The mills 109, 111 may each have a typical capacity of 5-10 t/h. A collector conveyor 112 is provided at the outlet of the fine mill 111 and joins the third secondary conveyor 108. On the path of the third secondary conveyor 108, provision is also made for sampling means 113 (for example a ladle), with which analysis means 114 may be fed.
Moreover, the third secondary conveyor 108 distributively feeds a first tertiary conveyor 115 and a second tertiary conveyor 117. The first tertiary conveyor 115 feeds a silo for storing waste 116. As for the second tertiary conveyor 117, it feeds a container 118, at the outlet of which a return conveyor 119 feeds the storage means 103. An air decontamination system 120 is set up at the coarse mill 109 and the fine mill 111 and feeds a sleeve filter 121, which may have a typical capacity of 5,000 Nm3/h. The sleeve filter 121 is connected to a chimney 123 as well as to a fines recovery line 122, which feeds the silo for storing waste 116. It is obvious that one skilled in the art will be able to adapt the thereby described means to the needs of the installation, for example by varying the number or the type of mills or the capacity of the different mills used.
Referring now to
A calcinated waste recovery line 203 is provided at the outlet of the pyrolysis ovens 202a, 202b, 202c and feeds two silos for storing calcinated waste 204a, 204b. The number of these storage silos may be varied depending on the needs of each installation. The calcinated waste recovery line 203 may be a jacketed conveyor provided with water cooling means. At the outlet of each silo for storing calcinated waste 204a, 204b a respective primary materials feed conduit 205a, 205b is provided (both of these conduits forming together the primary materials feed line 6).
Moreover, at the outlet of each pyrolysis oven 202a, 202b, 202c, provision is made for a respective gas exhaust conduit 206a, 206b, 206c (the whole of these conduits corresponding to the aforementioned gas exhaust line 4). Each gas exhaust gas conduit 206a, 206b, 206c feeds a respective post-combustion chamber 207a, 207b, 207c. A typical example of the volume of the post-combustion chamber 207a, 207b, 207c is 15 m3. Each post-combustion chamber 207a, 207b, 207c is further fed by a respective air intake conduit 208a, 208b, 208c.
A burnt gases collecting conduit 209 connects the outlet of the post-combustion chambers 207a, 207b, 207c to the inlet of a vertical cooling chamber 210. A water coolant feed line 211 is also provided at the inlet of the cooling chamber 210. For example, spraying ramps located in the high portion of the chamber may be provided. A pre-cooled gases recovery conduit 212 is provided at the outlet of the cooling chamber 210, and feeds a sleeve filter 214. An air intake conduit 213 is connected to the pre-cooled gases recovery conduit 212. The sleeve filter 214 may for example have a capacity of 4,000 Nm3/h. A fines withdrawal conduit 215 on the one hand and a stripped gases withdrawal conduit 216 are connected to the outlet of the sleeve filter 214. The stripped gases withdrawal conduit 216 feeds a chimney 217.
Now referring to
A respective selective post-dissolution emptying line 305a, 305b is provided at the outlet of each selective dissolution container 301a, 301b which feeds a respective press filter 306a, 306b. A system for collecting solids 307 is placed at the outlet of the press filters 306a, 306b and feeds a drying oven 308, at the outlet of which is found the pre-treated material conduit 309 (corresponding to reference 14 in
The other major component of this portion of the installation is the electrodeposition unit 314. The electrodeposition unit 314 comprises a certain number of electrolysis tanks 315a, 315b, 315c, 315d, 315e, the number of tanks (five in this example) being adaptable to the production needs. Each electrolysis tank 315a, 315b, 315c, 315d, 315e comprises a certain number of electrolysis cells depending on the production needs, for example eight in the present example. As an example, each electrolysis cell may have a useful volume of 4 m3 and contain 31 stainless steel cathodes and 30 lead/calcium anodes with a useful surface area of 1 m2 per face. The electrolysis tanks 315a, 315b, 315c, 315d, 315e are fed in parallel by a rich electrolyte transfer conduit 313 connected to the outlet of the tank for storing rich electrolyte 312. The electrodeposition unit 314 is completed by a system for stripping the cathodes 316.
At the outlet of the electrodeposition unit 314, a depleted electrolyte recycling line 317 is provided, which feeds a first tank for storing depleted electrolyte 318 (for example with a capacity of 60 m3) and a second tank for storing depleted electrolyte 319 (for example with a capacity of 25 m3). The first tank for storing depleted electrolyte 318 is the source for feeding the depleted electrolyte feed conduit 303. The second tank for storing depleted electrolyte 319 feeds a first stripping reactor 320 (for example with a capacity of 15 m3.). This first stripping reactor 320 is also fed by a lime feed line 321. At the outlet of the first stripping reactor 320 a first pulp withdrawal line 323 is connected, which feeds an additional press filter 324.
Moreover, the fines withdrawal conduit 215 described in connection with
A gas decontamination system 330 passes through the whole of the storage tanks 312, 318, 319 of the selective dissolution containers 301a, 301b and of the stripping reactors 320, 325 and feeds a washing tower 331. The storage tanks 312, 318, 319, the stripping reactors 320, 325 and the washing tower 331 may be in epoxy resin/fiber glass with a standard thickness. The washing tower 331 comprises at the outlet an additional acid juices withdrawal line 332, which may in return feed the tank for storing depleted electrolyte 318. The washing tower 331 may have a capacity of 5 m3, be provided with standard lining and operate with water.
Again referring generally to
Downstream from the lead covering containers 15, 18, means are found for casting anodes 25, which notably comprise a kettle of the type described in
The last major portion of the present installation relates to refining and notably encompasses references 27, 33, 34 of
An electrolysis decontamination system 507 provides collection of the gases at the Betts electrolysis 501 and at the Betts reactor 503, and their transfer towards a washing tower 508, with a typical volume of 5 m3. A washing juice collecting line 509 provides a return towards the second tank for storing depleted electrolyte 319 of
The Betts electrolysis unit 501 is moreover provided with means for stripping cathodes 510. The means for stripping cathodes 510 provide a cathode feed line 511 which itself feeds a kettle 512 providing melting of the cathodes. The kettle 512 is of the type described in
The Betts electrolysis unit 501 is moreover provided with means for scraping anode stubs 514 which feed an anodic sludges collecting line 516 (the whole forming an example of anodic sludge recovery means 28). This anodic sludges collecting line 516 feeds a unit for treating anodic sludges 517 which may comprise washing means, weighing means and/or means for storing in a safe. A washed sludges transfer line 518 connects the unit for treating anodic sludges 517 to an oxidation oven (with a power of 800 kW, 1 ton capacity, for example), which also receives at the entry an air or oxygen intake line 519. At the outlet of the oxidation oven 520, an ingot collecting line 521 may ensure return to the safe storing means of the anodic sludge treatment unit 516. The litharge feed line 505 is also connected to the outlet of the oxidation oven 520. A fumes collecting line 522 is also provided at the oxidation oven 520, it may be connected towards the same filtration system as the one provided at the lead covering containers.
Method For Treating Waste
An exemplary method for treating waste containing precious metals is described hereinbelow, in the case when the waste is used electronic cards. In this example, the method comprises 5 main phases:
milling;
pyrolysis;
copper extraction by selective dissolution (leaching);
lead covering; and
refining.
This example corresponds to the use of the treatment installation described above in connection with
The electronic cards are received at the means for receiving waste 101. The electronic cards arrive at the entry of the installation by batches (containers, big bags, barrels), which are weighed at the weighing means 102, labelled, recorded and stored at the storage means 103. The cards may arrive in three main forms:
1) entire cards requiring double milling before any treatment;
2) pre-milled cards requiring simple milling before any treatment, and
3) properly milled cards with a size less than 5 mm, not requiring any additional milling before treatment.
This is why with the transport system described above, the cards are directed according to their nature: either successively towards the coarse mill 109 and then the fine mill 111 (case 1 above); or directly to the fine mill 111 (case 2 above); or directly to the silo for storing waste 116 (case 3 above). The coarse mill 109 performs milling or grinding of the waste reducing them to a size of less than 25 mm, while the fine mill 111 performs milling or grinding of the waste reducing them to the required size of less than 5 mm. Moreover, before entering the properly milled cards into the silo for storing waste 116, the latter undergo automated sampling at the sampling means 113, which periodically interrupt the flow of cards. For example, 300 kg of sample per 24 t batch may be sampled. The sample is then analyzed by the analyzing means 114 after quartering in the laboratory in order to achieve a final sample mass of 4-5 kg. It is preferred to only treat a given batch of waste when the result of the analysis is known, in order to adapt the treatment parameters. This is why, before the sample analysis is carried out, the cards return via the return conveyor 119 to the storage means 103.
The premises of the mills 109, 111 are decontaminated, and the fines suspended in air are recovered and re-injected into the silo for storing waste 116. The milled electronic cards are then extracted at the base of the silo for storing waste 116 and feed the hoppers located above the entry of each of the three pyrolysis ovens 202a, 202b, 202c. The bulk density of the product is 0.7 at the entry of the ovens. Pyrolysis is useful for degrading and removing the organic materials contained in the cards. This is a controlled combustion of carbonaceous chains, which is carried out while maintaining the metals of the waste in the metallic state.
The dwelling time in the ovens may be comprised between 20 and 90 min and is preferentially 30 min. The operating temperature may be comprised between 350 and 550° C. and preferably have the value of about 400° C. By controlling the temperatures, the negative pressure and the screw velocity, it is possible to keep the operation under control. Each oven typically has a treatment capacity of 1 t/h. Pyrolysis gases rich in phenolic compounds emerge at 400° C. from each oven.
The calcinated cards exiting each oven are cooled on the calcinated waste recovery line 203 (jacketed conveyor) and are then stored in the silos 204a, 204b feeding the copper leaching. The product appears with a black appearance due to residual carbon from pyrolysis of the plastics. It has a density of about 0.5.
The pyrolysis gases from each oven are burnt at a high temperature in the post-combustion chamber 207a, 207b, 207c (dwelling time of 2 s) in order to destroy all the carbonaceous molecules and possible dioxins and furans. Thus, quasi the whole of the carbon chains is recovered in the form of energy which may be recovered and used in the actual process. The oxidizing air is preheated to 400° C. in order to provide proper inflammation of the gases. A controlled air supplement is required for regulating the chamber exit temperature to 1,100° C. and avoiding formation of NOR. A supplemental 800 kW burner ensures that the temperature is sufficient for the combustion to occur, notably during the transient phases. Continuous control of the entry and exit post-combustion temperatures is achieved and the incoming dilution air may be regulated.
The gases at 1,100° C. arrive in the cooling chamber 210 in order to be subjected therein to water quenching. The coolant water is fed at a flow rate of 10 m3/hour. The water is entirely transformed into steam by absorbing a considerable portion of the energy of the gases. The cooled gases exit the chamber at about 200° C. By controlling the exit temperature, it is possible to regulate the injected water flow rate.
These gases are then finely cooled with air to about 150° C. before entering the sleeve filter 214. The sleeves retain the fine solid particles, notably containing halides. Their purge is carried out in the following sector of copper leaching. The totally stripped gases are rejected into the chimney 217. Continuous monitoring is carried out therein (analysis of the gases, dust levels . . . ). As regards extraction of copper, the latter is carried out by means of a sequence of hydrometallurgical operations: leaching in an oxidizing acid medium, filtration of the residue, electrodeposition of the copper requiring the use of several reactors, storage tanks, filters, pumps and a set of electrolysis cells and of current generators.
The daily throughput of calcinated cards coming from the silos 204a, 204b containing 12 t of copper is treated in selective dissolution containers 301a, 301b (closed reactors) in 11 leachings, each lasting 4 hours. The operation is as follows: transfer to the pump of 15 m3 of copper-depleted and acid-rich electrolyte (at 85° C.) from the tank for storing depleted electrolyte 318. One then proceeds by introducing 4.8 t of calcinated cards with fine oxygen bubbling at the bottom of the reactor. The depleted electrolyte is a solution containing sulphuric acid (50-200 g/L, preferably about 100 g/L) and soluble iron as iron sulphate (5-20 g/L, preferably about 10 g/L) which has to be maintained in the form of Fe3+ (with oxygen) for efficiently etching the copper.
Maintaining the temperature is provided by injecting fresh steam. Finally, the contents of the reactor are filtered on the press filter 306a or 306b or both. The copper-rich juices are transferred towards the tank for storing rich electrolyte 312 feeding the electrolysis cells.
The electrolyte is enriched with iron and nickel which are dissolved at the same time as copper. A daily purge has to be carried out on the depleted electrolyte exiting the electrolysis cells. It is sent to the second tank for storing depleted electrolyte 319 and its treatment is carried out in the first stripping reactor 320 twice daily. These very acid juices containing iron, nickel and a little copper are treated with lime up to a pH of 8.5. Calcium sulphate precipitates carrying along metal hydroxides. This pulp is filtered on the additional press filter 324. The obtained residue (10-15 t/d) is placed in a landfill site. The juices are recycled to the first tank for storing depleted electrolyte 318. The fines of the filter of the pyrolysis are treated in the second stripping reactor 325 every 2 days in the presence of water and some lime at pH 9. The halides (mainly chlorides and bromides) pass into the solution. The pulp is filtered on the additional press filter 324: the residue (500 kg) is recycled to the selective dissolution containers 301a, 301b and the juices (3 m3) enriched in halides are stored in a tank body for subsequent treatment.
The whole of the reactors, storage tanks, filters, are decontaminated and the vapors and droplets are absorbed by the washing tower 331. The obtained acid juices are regularly purged and recycled to the tank for storing depleted electrolyte 318.
In the process, the depleted electrolyte is raised to 85° C. and maintained at this temperature by means of a coil fed with steam. The rich electrolyte is cooled to 50° C. by a coil fed with cold water. This cold water may then be used in the chamber for vaporizing the hot gases from the post-combustion of the pyrolysis.
The wet solid residues (40 t/d) stemming from the press filter 306a and 306b are rich in precious metals. They are dried in the drying oven 308. They are powdery and have a black color, the glass fibers which are the main compound thereof being broken during the stirring in the etching tank. The solid and liquid flows are regularly sampled and analyzed.
As regards the electrodeposition step, the flow stemming from the rectifiers, passes in series from electrolysis cell to electrolysis cell and in parallel at the electrodes of each cell. Current density may be from 50 to 400 A/m2, preferably about 200 A/m2 and the temperature of the electrolyte may be from 20 to 80° C., preferably from 45 to 50° C. The concentration of ferric ions is maintained as low as possible, and in any case at a level less than 10 g/L. When the total iron concentration reaches a value of 10-30 g/L, a portion of the electrolyte is purified by precipitation of iron and filtration of the precipitate.
The rich electrolyte coming from the tank for storing rich electrolyte 312 is sent into the 1st row of eight cells. The cells are positioned as a cascade so as to allow circulation of the electrolyte and a pump sends back the juices from the last cell to the first. The circulating flow rate is of the order of 15 m3/h. The electrolyte takes 24 hrs for being depleted in copper which is deposited on the cathodes. By adding a surfactant, a fine and regular copper deposit may be obtained. The depleted electrolyte is pumped towards the tank for storing depleted electrolyte 318. The cells are then again filled with rich electrolyte. Each row may be emptied and then filled with electrolyte every 4.5-5 hrs.
Every 5 days, a row of cells is stripped. The recovered copper cathodes (12 t) are rinsed with water and the copper is separated from its stainless steel support by a suitable mechanical tool. The obtained copper is sampled and stored. It should further be noted that the whole of the solid and liquid flows are advantageously regularly sampled and analyzed.
The step of selective copper extraction is important when the starting material contains a large copper proportion. Indeed, copper is able to form stable compounds which are insoluble in liquid lead, said compounds often containing precious metals. This is why it is necessary to get rid of the largest amount possible of copper before starting with the following lead covering and refining steps, otherwise a large amount of precious metals would be lost in said stable compounds.
Moreover, the selective copper extraction step enables quasi the whole of the copper to be selectively extracted as a commercial product (pure copper cathodes), which may be re-melted as ingots. The metals dissolved in the electrolyte (iron, aluminium, nickel) may advantageously be removed during periodic operations for regenerating the electrolyte. It should be noted that electrodeposition of copper may be replaced with an operation of copper sulphate crystallisation, which is a commercial product.
According to the type of waste, it is possible to do without either of the pyrolysis and copper extraction steps by selective dissolution or both. This is the case for example when the waste consists in catalytic mufflers. In this case, the carbon chain and copper contents do not justify the presence of both of these steps, and the milled used catalytic mufflers are directly subjected to the lead covering step.
The lead covering step comprises contacting the pre-treated materials (i.e. after milling, optional pyrolysis, optional copper extraction) with a molten lead-based composition in the lead covering container 15. The molten lead-based composition comprises lead in majority and may comprise from 0 to 50% tin, preferably less than 20% tin. This composition is in the liquid state. It is used as a collector and extractor of precious metals, which are found solubilized in a non-oxidized form. The lead and optionally the tin of this composition partly stem from added metals contained in electronic cards, and partly from metals recovered subsequently.
Dissolution is carried out in the following way: stirring is started and it generates a vortex of molten lead in the kettle. The feeder 407 pours the materials to be covered with lead into the core of the vortex. The operation lasts for about 15 minutes. The temperature may then be comprised between 350 and 550° C. and preferably is about 500° C.
A phase of skimming or phase of separating the elements without any affinity for lead is started subsequently. In this phase, stirring is stopped, and the inert portions with their precious metals (ceramics, glass fibers, ferrites . . . ) having been washed away move up to the surface where they float. The skimming machine 405 is then started and with it the supernatant materials may be recovered. When these supernatant materials have been removed from the lead bath, the operation is repeated. Skimming may be carried out at a temperature comprised between 250 and 450° C., for example of about 270° C. for a lead-tin alloy with 30% by weight of tin.
The supernatant materials may again be treated in the same way in the additional lead covering container 18. Indeed, a small amount of lead (and of precious metals) is carried away with the supernatant materials during skimming, and it is therefore useful to repeat the operation in a second kettle in order to avoid loss of precious metals. The inert materials collected upon skimming at the additional lead covering material 18 are sent to a landfill as ultimate waste after having been optionally sampled and analyzed. The molten lead-based composition contained in the additional lead covering container 18 has a small concentration of precious metals (less than 100 g per ton) and it is sent back by pumping towards the lead covering container 15.
The lead covering phase may last for several days. It is considered that it is completed when the precious metal content in the molten lead-based composition reaches a threshold value, for example located between 2 and 4 kg per ton of lead. One may then proceed with an optional decoppering operation, consisting of adding sulphur in the lead-based composition vortex, in order to form copper mattes which are sent back into a selective dissolution container 301a, 301b for extracting copper. Next, the molten lead-based composition with the solubilized precious metals is sent to a storage kettle from which this composition is cast into anodes (this forms the means for casting anodes 25).
With the following Betts refining step, the precious metals contained in the thereby cast anodes may be released. During this step, lead and tin are removed from the anodes by electrolysis in a fluorosilicic medium, which is known as the Betts process, at the Betts electrolysis unit 501. The electrolysis cells are fed with an electrolyte for example containing about 90 g/L of lead and 80 g/L of free acid. This electrolyte is prepared by solubilizing litharge (PbO) in fluorosilicic acid at the Betts reactor 503. Electrolysis may be carried out for example at a current density of 350 A/m2 and at an electrolyte temperature of 40° C. During the electrolysis, the solid lead and tin of the anodes are dissolved in the electrolyte, while the deposits of lead and tin build up on the cathodes. By adding surfactant in the electrolyte, it is possible to make this deposit fine and smooth. The precious metals themselves remain at the anodes.
In this configuration, the electrolyte does not concentrate impurities or only very little. A total purge may only be carried out once or twice a year. In this case, the electrolyte is sent to the level of the second tank for storing depleted electrolyte 319 for a lime treatment in the first stripping reactor 320, and a new electrolyte is prepared. A partial purge may be required for lowering the lead content with sulphuric acid in the Betts reactor 503. The reactor 503 and the electrolysis cells are decontaminated and the gas effluents are sent to the washing tower 508. The washing juices are treated in the first stripping reactor 320 via the second tank for storing depleted electrolyte 319.
In the example described herein, the two rows of five cells contain 60×5=300 anodes each of 450 kg. Six days are required for consuming 80% of the anodes which therefore release their precious metals (432 kg) as anodic sludges, i.e. dissolved anode residues. The anodic sludges may, depending on the cases, fall into dust in baskets or have an adhering cell structure. Preferably electrolysis is interrupted before complete dissolution of the anodes. Anodic sludges may thereby be recovered by scraping.
The produced Ph/Sn cathodes represent 108 t for these six days. Two strippings of cathodes are advantageously performed every three days and two scrapings of the anodes at the same time. The cathodes are recycled towards the kettle 512 in order to produce commercial lead-tin in ingots and/or for feeding the additional lead covering container 18 with a molten lead-based composition.
The anodic sludges are washed, weighed and stored in a safe. Advantageously, the anodic sludges are melted once to twice a week in the oxidation oven 520 at a temperature of 1,000° C. This step of melting anodic sludges, in the presence of oxygen gas (or air) enables at least a portion of the lead and of the tin to be oxidized, which were still contained in the anodic sludges. Litharge (PbO) forms at the surface: it is cast into plates and it may dope the electrolyte with lead when necessary. The fumes from the oven are channeled towards the lead covering filter. The liquid precious alloy may be cast into ingots (25 kg) which are stored batchwise in a safe. All the ingots are sampled and then weighed before being marketed.
The method described above is designed in order to limit as much as possible the losses of precious metals at all stages of the method.
with the fines recovery line 122, the fine fragments of milled waste which pass into ambient air at the milling, may again be introduced into the system;
with the fines withdrawal conduit 215, the metal fraction carried away with the carbonaceous gas during pyrolysis may be recovered;
with the washed fines withdrawal line 327, it is possible to re-inject into the main circuit the fragments of materials to be treated carried away with the electrolyte(s);
by using an additional lead covering container 18 in addition to the lead covering container 15, precious metals inadvertently carried away with the inert materials during the main lead covering may be recovered;
besides, the copper which was not extracted during the selective extraction step by dissolution is recovered as copper mattes at the lead covering container 15 and is re-injected into the main circuit.
Thus, with the method, it is possible to recover in fine more than 90 or even 95% by weight, preferably more than 99% by weight, advantageously more than 99.9% by weight of the precious metals initially contained in the waste.
Tables 1 and 2 below give an estimation of the chemical composition of the products during the different steps of the treatment method, in the case when the waste is typical used electronic cards.
| Number | Date | Country | Kind |
|---|---|---|---|
| 06 09 219* | Oct 2006 | FR | national |
This application is a National Phase Entry of International Application No. PCT/FR2007/001728, filed Oct. 19, 2007, which claims priority to French Patent Application No. 0609219, filed Oct. 20, 2006, both of which are incorporated herein by reference.
| Filing Document | Filing Date | Country | Kind | 371c Date |
|---|---|---|---|---|
| PCT/FR07/01728 | 10/19/2007 | WO | 00 | 4/17/2009 |
