|Title of Invention||
A METHOD FOR THE PRECIPITATION OF SILICA IN CONNECTION WITH ZINC ORE LEACHING
|Abstract||The invention relates a method for the leaching of oxidized ores and in particular zinc ores. The ore, wherein the valuable metals are at least partially in silicate form, are routed to an acidic leaching stage in conditions where the silicate decomposes and the valuable metal ion comes into the solution. During leaching the silicate ion first dissolves, but simultaneously decomposes and is precipitated as silica.|
|Full Text||FORM - 2 THE PATENTS ACT,1970
(39 of 1970)
(Section 10, rule 13)
A METHOD FOR THE PRECIPITATION OF SILICA IN CONNECTION WITH ZINC ORE LEACHING
of Riihitontuntie 7, FIN-02200 Espoo, Finland a Finnish public limited Company
THE FOLLOWING SPECIFICATION PARTICULARLY DESCRIBES
THE NATURE OF THIS INVENTION AND THE MANNER
IN WHICH IT IS TO BE PERFORMED:-
This invention relates to a method for the leaching of oxidized ores and in particular zinc ores. The ore, wherein the valuable metals are at least partially in silicate form, are routed to an acidic leaching stage in conditions where the silicate decomposes and the valuable metal ion comes into the solution. During leaching the silicate ion first dissolves, but simultaneously decomposes and is precipitated as silica.
The majority of the world"s reserves of zinc occur in so called oxidized ore, where the zinc is bound to silicates and carbonates. Such are for instance smithsonite (ZnCO3) and willemite (2ZnOSiO2) and mixtures of these minerals. Zinc-containing siliceous slags also exist, which originate mainly from the production of lead. Thus in the text the terms ore or raw material are also used to mean other zino-containing siliceous raw materials than actual ores. Industrially these ores are utilized both pyrometallurgically and hydrometallurgically. Hydrometallurgical utilization occurs by leaching. The carbonate portion does not cause any major difficulties in leaching, but on the other hand the siliceous fraction is hard to control. Various processes have been proposed to overcome these difficulties,
US Patent document 4,148,862 (no Indian equivalent application available) describes how the problem of siliceous zinc ores relates to the precipitation of silica, SiO2 and especially the morphology of silica. In industrial operation, the precipitation rate in particular is problematic. In the US publication in question, there is a diagram that illustrates the dependence of the stability of silica gel in an aqueous solution on the pH value. This shows that the precipitation rate of silica is at its best in a very acidic solution and again in a pH range of 3 - 5, but that a pH of 2 is very unfavourable for precipitation. It is clear from the patent text that the problems arising in connection with a siliceous zinc ore are quite different from those in normal zinc processes, in which the raw material is a sulphidic concentrate.
At least three ways have been proposed to combat the problem of a siliceous zinc concentrate. The first of these is described in for instance US Patent 3,656,941 (no Indian equivalent application available), where the leaching of a siliceous material performed in acidic conditions is rapid, so that the decomposed silica remains in solution. The solution is transferred to a second stage, where the SiO2 is precipitated. Precipitation occurs by neutralizing the solution with a calcine, limestone, lime or other suitable neutralizing agent.
When neutralization takes place as a continuous process, as described in US Patent 3,656,941 (no Indian equivalent application available), the solution can be filtered well even though the dissolved silica is up to 50 g/1. On the other hand however, it is known that the solution is unstable with regard to silica and sooner or later the silica will start to precipitate from the solution as a unfilterable gel. Therefore such a process is risky on industrial scale since one must always be prepared for stoppages that over time lead to the uncontrollable precipitation of silica.
Another treatment method of siliceous zinc ore is described in US Patent 3,954,937 (no Indian equivalent application available), where in the siliceous material leaching the pH value of the solution is lowered gradually so that at the end of the leach it is about 1.5 i.e. the H2SO4 content is about 1.5-15 g/l. The reduction of the pH value is carried out so slowly that the silica is able to precipitate. In a continuous process the siliceous material is brought to the first reactor and the acid is added to each reactor in the direction of the slurry flow. The drawback of this method is that the process requires the extremely careful regulation of each reactor. If too much acid goes into one reactor, silica is precipitated uncontrollably and a precipitate difficult to filter is obtained.
The above-mentioned US Patent 4,148,862(no Indian equivalent application available) describes a third method, where the leaching of a siliceous material is performed in a single reactor and the
pH is held constant (maximum 2.5) throughout the duration of the leach. Residence time is prolonged until the silica is able to precipitate. Although it is not stated in detail in the patent, it is evident that the solution coming from this acidic leach must be neutralized after filtration before it can be transferred forward. In order for the amount of neutralizing agent to be kept to the minimum, it is of course worth keeping the pH in the leaching and precipitation stage of the siliceous material as high as possible.
Now a method has been developed, whereby it has been made possible to eliminate the drawbacks of the above-mentioned processes. This method enables the treatment of silicate-containing zinc ores so that the highest possible zinc yield and a leaching residue that settles and filters well are obtained. It has been shown, surprisingly, that it is possible to obtain a leaching residue where the silica has excellent filtration properties. This result is achieved if an acidic leach, where the silicate is leached and precipitated as silica, is carried out so that the highest acid content (lowest pH) is at the start of the leaching and the lowest acid content (highest pH) is at the end of the leaching. In a continuous process this occurs so that the silicate-containing material is taken to a leaching stage containing several reactors, where the highest acid content is in the first reactor and the lowest acid content in the last reactor. The reactions are made to reach the end before the slurry is taken to the separation stage. The final pH of the leach is around the value of 1.5 - 2. The majority of the leaching occurs thus in a high acid content, where the silicate ion decomposes and dissolves and the leaching yield is excellent, and the silica precipitates quickly. Only at the end of the leach is the pH raised to the region where both dissolving and silica precipitation are slower. At this stage there are however already present SiO2 nuclei formed in the early stage of the leaching, and these facilitate the precipitation of the silicate ion as silica. The essential features of the invention will be made apparent in the attached claims.
The invention is explained in more detail using the appended drawings, where Figure 1 shows a flow sheet of one embodiment of the invention, and Figure 2 shows a flow sheet of another embodiment of the invention.
Flow sheet 1 presents a simplified illustration of a continuous leaching process of a silicate-containing raw material. This shows a method for the treatment of a silicate-containing ore, but obviously it can be adapted for the leaching of other silicate-containing materials. The ore 1 to be treated is first taken to a neutral leaching stage 2, where the ore is leached into a dilute sulfuric acid-zinc sulfate solution. Leaching occurs in atmospheric conditions. At this stage the easily soluble zinc compounds in the ore decompose and form zinc sulfate. At the end of the leach the pH of the solution should be sufficiently high, at a pH of about 5, for the further treatment of the zinc sulfate solution. If this value is not reached otherwise, the solution is neutralized for example with calcine (ZnO) or lime.
After the neutral leach the slurry is taken to separation of solution and solids. Separation occurs in a neutral leach thickener 3, from which a zinc sulfate solution is obtained as an overflow 4, which is taken via solution purification to zinc electrolysis for the production of elemental zinc (not shown in the drawing). A thickener underflow 5 contains silicates, which do not dissolve in neutral leach conditions. The underflow is routed to an acidic leaching stage 6. The acidic leach occurs using a return acid 7 from electrolysis, which contains 150 - 220 g/l of free sulfuric acid. The acidic leaching stage is carried out in a reactor series comprising several reactors. In the drawing there are three reactors, but in practice the number of reactors is selected so that the desired reactions proceed in them right to the end. The leaching stage is performed in atmospheric conditions at a temperature of about 60 -100°C.
The return acid is largely all fed to the first reactor 8, wherein the pH in the first reactor is the lowest and the acid content therefore the highest. The
silicate-containing precipitate 5 is routed to all the reactors in the reactor series so that the pH of the leaching stage is raised gradually towards the final reactor 9 of the stage, where the pH is around a range of 1.5 - 2. At a higher pH the leaching yield of zinc weakens. The precipitate 5 can be fed for instance to each reactor in equal amounts, whereby the pH of the slurry rises step by step as the leaching progresses. This is technically the easiest way to feed in the precipitete, but obviously it can be done in other ways as the need arises. The leaching time of silicate-containing precipitate in the acidic leaching stage is between 3 - 15 h.
From the final acidic leaching reactor the slurry is taken to solution and solids separation 10 i.e. thickening and filtration. The underflow obtained from the post-acidic leaching filtration i.e. the leaching residue 11 includes the silicates contained in the ore as silica. The leaching residue is removed from the leaching circuit, but if needed this may also be recirculated back to the first reactor 8 of the acidic leaching stage. The overflow 12 from the filtration stage 10 is a zinc sulfate-sulfuric acid solution, with a sulfuric acid content of 5-15 g/l. The overflow is led to the neutral leaching stage 2 for neutralization by the ore 1 fed there.
The flow sheet in Figure 2 is similar to that shown in Figure 1 except that in this case the precipitate 5 is no longer taken to the final reactor 9 of the acidic leaching stage 6. If the reactions occurring in the leaching stages continue in the thickener, they will disturb the operation of the thickener. When solids are no longer taken to the final reactor, the reactions have time to proceed to completion there and there are no unreacted solids left in the slurry to be taken to the thickener. If for example the dissolving of carbonates continues even in the thickener, the gaseous carbon dioxide (CO2) released could form a layer of toam on the surface of the thickener and prevent the separation of solids. The flow sheet also shows that the reactor slurry may be recirculated in the acidic leaching stage so that the slurry from some of the reactors at the final end of the flow direction is recirculated to the first
reactor of the stage. In the case shown in Figure 2 the slurry is recirculated from the solids separation stage 10 to the first reactor 8 of the acidic leach.
One way to implement the method of this invention is to carry out leaching in batches, whereby in principle only one leaching reactor is needed, which is filled first with return acid, and silicate-containing slurry is fed into the return acid so that the pH of the slurry rises step by step. In this way the acid gradient is as even as possible, but in industrial implementations the continuous process is generally the most economical method.
The process can also be utilized in such cases where a zinc liquid-liquid extraction follows the leaching stages described above. The use of extraction becomes relevant if the Zn content of the ore is so low that the water balance of the normal Zn process is difficult to control, because the high amounts of wash water dilute the solution too much. In such a case it is economically advantageous to operate at a low temperature, even at room temperature, because a high temperature in extraction is harmful, in contrast to the normal Zn process.
The method of the invention is further described with the aid of the following examples:
A test was performed as a batch, where silicate-containing ore was added evenly to a solution of return acid. The ore contained the following: SiO2 15.5%, Zn 27% and Pb 6.5%. The H2SO4 content of the return acid solution was 195 g/l and the Zn content 52 g/l. Ore was added over a period of 7 h, after which the pH of the solution was 1.5 and its H2SO4 content 7.8 g/l. The amount of ore added was 345 g per liter of return acid. The temperature of the batch was 90°C.
After leaching, 120 mg of flocculants per kg of leaching residue were added to the slurry and settled using a rake. The underflow from the settling test was filtered and washed in a membrane pressure filter. The filtration capacity was 98 kg/m2h and the solids content of the underflow was 550 g/l. The analysis of the leaching residue was as follows: Zn 1.3%, Pb 13.6% and SiO2 31.9%. The SiO2 content of the filtration overflow, a zinc sulfate solution, was 460 mg/l. The leaching yield of zinc was 97.3%.
A test was performed on a continuous basis using the apparatus shown in Figure 1. The composition of the return acid was the same as in example 1. The return acid was fed into the first reactor of the acidic leach at a rate of 1 l/h. Ore was fed into the neutral leaching stage at a rate of 360 g/h. The neutral leaching stage underflow was divided between the acidic leaching stage reactors so that seen from the direction of the flow the H2SO4 content in the various reactors came to 130, 70 and 10 g/l and these corresponded to pH levels of 0.4, 0.8 and 1.5. There was a residence time of 3h in each reactor and the temperature was 90°C.
The results of the tests were as follows: Zn yield was in the region of 97.8 -98.4%. Thickening was carried out after the acidic leaching, which produced an underflow with a solids content of 540 - 680 g/l. A pressure filtration capacity of 96-128 kg/m2h was achieved. Filtration tests were made without dilution, and so the specific weights of the solution were between 1.38 - 1.41 g/cm3. The SiO2 content of the zinc sulfate solution obtained as the filtration overflow was 350 - 550 mg/l.
In addition several vacuum filtration tests were made, which gave a filtration capacity of 60-100 kg/m2h, when thickening was made with a diluted solution. This represents the use of a combination washing-thickener i.e. the acidic leaching slurry was washed before thickening. The specific weight of
the solution going to thickening was about 1.25 g/cm3 and the moisture content of the precipitate obtained was about 35%.
The examples show that the results are of the same order as in a zinc process, where a calcine prepared from a sulphidic concentrate with a silicate content of the order of 1 - 1.5% is used.
When the results of the above-mentioned examples are compared with those for example obtained using the US patent 4,148,862 method, it can be seen that the underflow obtained from filtration after acidic leaching has a solids content of around 500 g/l, whereas in the above-mentioned US patent it is only in the region of 200 g/l. In the same US patent the moisture content of the precipitate is over 60% (wet cake 33.7 g and dry cake 12.3 g), whereas in the method according to the invention it is in the region of 35%. The results are very valuable for industrial operation, since in the method of the invention the solution flows are reduced and the amount of wash water is decreased compared with the methods of the prior art. This is very important when using a raw material that produces 3-4 times more leaching residue to be washed per tonne of zinc formed than a normal zinc sulfide process
The examples given above are only various ways to make use of the method of the invention, and clearly there are aiso other ways to exploit it. The method is not tightly bound to the residence times given in the examples.
1. A method of leaching zinc-bearing siliceous raw materials for the recovery of zinc, characterized in that the siliceous material is taken to an acidic leaching (6) occurring with a sulfuric acid solution (7)), where the highest acid content and thus the lowest pH is at the start of the leaching, whereby the silicates in the material decompose and dissolve and precipitate as silica, and the acid content is lowered as the leaching proceeds so that the pH of the solution at the end of the leach is around in the region of 1.5 - 2.
2. A method according to claim 1, characterized in that the leaching temperature of the siliceous material is between 60 - 100°C.
3. A method according to claim 1, characterized in that there are SiO2 nuclei in the solution at the end of the acidic leach.
4. A method according to claim 1, characterized in that the zinc-containing siliceous raw material is leached in a continuous process, whereby the raw material (1) is fed first to a neutral leaching stage (2), where the easily soluble zinc compounds in the raw material decompose, and where a zinc sulfate solution (4) obtained after a separation (3) is taken to electrolysis and a precipitate (5) is taken to the acidic leaching stage (6), where leaching occurs in several reactors (8,9), in which the pH of the slurry is raised in the direction of the flow to be in the region of 1.5 - 2 in the final reactor (9).
5. A method according to claim 4, characterized in that the silicate-containing precipitate in the acidic leaching (6) is leached with an electrolysis return acid (7), so that the whole amount of acid is routed to the first reactor (8) in the reactor series and the precipitate is taken
to several reactors in order to raise the pH of the leaching stage gradually towards the final reactor (9).
6. A method according to claim 4, characterized in that an equal amount silicate-containing precipitate is taken to each reactor in the acidic leaching stage.
7. A method according to claim 4, characterized in that no silicate-containing precipitate is taken to the final reactor in the acidic leaching stage.
8. A method according to claim 4, characterized in that the slurry from the acidic leaching stage is recirculated from some of the reactors at the end of the flow to the first reactor in the stage.
9. A method according to claim 4, characterized in that the leaching time of the silicate-containing precipitate is between 3 - 15 h.
10. A method according to claim 4, characterized in that the slurry from the final reactor in the acidic leaching stage is taken to a solution and solids separation (10), from where the solution obtained (12) is taken to the silicate-containing raw material neutral leaching (2) and a precipitate containing silica (11) is removed from the leaching circuit.
11. A method according to claim 4, characterized in that the precipitate (11) coming from the acidic leaching stage separation is recirculated to the first reactor of the acidic leaching stage.
12. A method according to claim 1, characterized in that the silicate-
bearing raw material is leached on a batch principle, whereby all of
the silicate-bearing material is taken into the sulfuric acid-containing
solution in a way that the pH of the solution rises step by step.
13. A method according to claim 1, characterized in that the silicate-bearing material is zinc ore.
14. A method according to claim 1, characterized in that the silicate-bearing material is zinc-bearing slag.
DATED THIS 1ST DAY SEPTEMBER 2003.
OF R.K.DEWAN & COMPANY APPLICANT"S PATENT ATTORNEY
|Indian Patent Application Number||817/MUMNP/2003|
|PG Journal Number||46/2007|
|Date of Filing||01-Sep-2003|
|Name of Patentee||OUTOKUMPU OYJ|
|Applicant Address||RIIHITONTUNTIE 7, FIN-02200 ESPOO,|
|PCT International Classification Number||C22B 3/08|
|PCT International Application Number||PCT/FI02/00182|
|PCT International Filing date||2002-03-08|