AU734903B2

AU734903B2 - Metal recovery process

Info

Publication number: AU734903B2
Application number: AU28617/97A
Authority: AU
Inventors: Wayne Bevan Diconoski; Fook Sin Wong
Original assignee: MIM Holdings Ltd
Current assignee: MIM Holdings Ltd
Priority date: 1996-07-15
Filing date: 1997-07-14
Publication date: 2001-06-28
Anticipated expiration: 2017-07-14
Also published as: AU2861797A

Description

AUSTRALIA

Patents Act 1990 COMPLETE SPECIFICATION FOR A STANDARD PATENT Name of Applicant: Address for Service: M.I.M. HOLDINGS LIMITED CULLEN CO., Patent Trade Mark Attorneys, 240 Queen Street, Brisbane, Qld. 4000, Australia.

Invention Title: METAL RECOVERY PROCESS Details of Associated Provisional Application: No. P00992 The following statement is a full description of this invention, including the best method of performing it known to us.

This invention relates to a metal recovery process and particularly relates but not limited to a process for recovering zinc from a lead blast furnace slag or from any other type of slag, residue or metalliferous material which can contain iron, silica and zinc. The invention can also extend to the recovery of other metals such as copper from a copper converter slag, or nickel, cobalt and other metals The invention will be described using a lead blast furnace slag as a suitable material.

In an example, blast furnace slag from lead smelting operations contain an appreciable amount of zinc, typically between 11% 20%. A large lead smelting operation has an annual production of 250 thousand tonnes of slag and many operations also have a slag stockpile of several million tonnes.

Several attempts have been made to separate the zinc from the slag, but these processes have not been economical as they required large acid costs. Energy ooooo costs are also appreciable. For instance, an attempt has been made to remove zinc from lead slag by sulfation 0.

roasting with ferrous sulphate at 650C; and chlorination with salt at 1,000 0 C. A Zinc oxide fuming process is also known which requires a high energy input, making it 25 uneconomical to recover metals from a slag.

o• A known method for recovering zinc from a siliceous zinc ore is to leach the ore with aqueous sulphuric acid. The zinc dissolves into the sulphuric acid to give a zinc sulphate leach solution. Impurities 0:0. 30 are removed from the solution (usually by precipitation), and the solution is then subjected to solvent extraction followed by electrowinning to deposit the zinc from the electrowinning solution.

There are two main disadvantages with the use of the above leach process to recover metals such as zinc from slags. Firstly, as the amount of zinc in the slag is relatively low, and due to the presence of silica, a large amount of fairly expensive sulphuric acid is required. This increases the cost of the zinc recovery to a level which is uneconomic.

The second main disadvantage with acid leaching of the slag is that the slag can contain between 10% 20% silica. The silica also dissolves in the sulphuric acid, and readily forms an unfilterable gel which occludes the zinc sulphate.

The silica problem has been previously identified in the recovery of zinc from zinc silicate ores, the two main ores being willemite and hemimorphite.

These ores are primarily zinc silicate ores having only small amounts of impurities. Acid leaching of these ores causes the silica to gel or polymerise to make it virtually impossible to remove the zinc sulphate.

.Only by adopting extremely careful and rigid process parameters is it possible to precipitate the silica from solution at an acceptable rate and to minimise formation of the gel. U.S. patent 3,656,941 to Electrolytic Zinc Company of Australasia describes a 20 process for recovering zinc from zinc silicate ores. A similar difficulty arises with recovering copper from copper silicate ores such as crysacolla The process described in the above U.S. patent cannot be simply applied to recovering zinc from blast 25 furnace slags, as the slags, as well as containing silica, also contain a appreciable amounts of iron, calcium, and the zinc content is primarily in the form of zinc ferrite as well as a zinc silicate.

The present invention is directed to a process to more economically recover metals from a silica :i containing material such as a slag, the invention having developed a method by which silica can be readily precipitated, and optionally where the aqueous sulphuric acid required for the leach can be produced in an economical manner.

It is an object of the invention to provide a process which may overcome the abovementioned disadvantages or provide the public with a useful or commercial choice.

According to a first broad form of the invention there is provided a method for removing silica from a leach solution obtained by leaching a material comprising a metal and silica in an aqueous acid, said method comprising ensuring said leach solution comprises a level of iron in its ferric state to allow precipitation of the iron as goethite which also results in precipitation of the silica.

According to a second broad form of the invention there is provided a method for recovering a metal from a material comprising the metal and silica, said method comprising leaching the material with a leach solution comprising an aqueous acid to dissolve the metal and the silica, removing the silica by the method of the first broad form and separating the metal from the leaching solution.

The material to be leached can be a slag such as a lead blast furnace slag containing zinc, a copper 20 containing slag, and any other type of slag, residue or metalliferous material which contains the metal to be recovered and silica.

The at least one metal can be zinc, copper, nickel, cobalt or combinations thereof, or other 25 desirable metals which dissolve under the leach "":"conditions If the material has an appreciable iron content, the iron will be dissolved into the leach solution to provide the iron level. Iron can be added or removed to the solution to provide a suitable iron level.

The precipitation step should precipitate sufficient iron as goethite to cause precipitation of essentially all, or at least effectively all the silica.

If the process is used to recover zinc, it is preferred that essentially all the iron is also precipitated from solution, as iron will contaminate downstream processing steps, such as the solvent extraction step. If the process is used to recover copper, not all the iron needs to be precipitated as iron does not contaminate downstream copper processing steps.

The process manages to overcome the difficulty with removing silica from the leach solution in an economical manner by ensuring that the iron in the leach solution is precipitated as goethite which precipitates readily from solution and also causes the silica to precipitate readily from solution.

The process described above to produce sulphuric acid can be used to recover other metals from materials containing the metal and silica.

To ensure precipitation of dissolved iron in the leach solution as goethite, the leach solution may need to be adjusted to minimise precipitation of the iron in other forms, for instance as ferric hydroxide. To promote precipitation of iron as goethite, the leach solution may need to be adjusted to ensure that the iron in the solution is in the ferric form, the ferric concentration is less than about 2g per litre and that 20 the temperature of the leach solution is maintained above at least 70 0 C and has a pH of about 3.8 or higher.

Subsequent to leaching of the material, the leach solution can be transferred to at least one iron/silica precipitation stage, and it is preferred that 25 a number of precipitation stages are used for longer o residence times. The resultant iron and silica-free leach can then be filtered to remove any remaining solids, and subjected to solvent extraction and electrowinning to recover the zinc. The raffinate can be recycled to the leach solution if desired.

To promote production of the ferric iron in the precipitation stage, an oxidant such as oxygen or air can be blown into the leach solution.

The material (such as a slag) may be leached by sulphuric acid, and it is preferred that the sulphuric acid is formed by an economical process which produces sulphuric acid by reacting a mixture of sulphur dioxide and air with a ferrous catalyst.

It is preferred that the mixture has about 3% and preferably between 15% 16% sulphur dioxide, and this percentage is present in the off gas from copper smelters, lead smelters, or from burning sulphur.

A detailed description of the process will be described with reference to the accompanying flow diagram illustrated in Figure 1.

In an embodiment, blast furnace slag is used as the material for leaching. The blast furnace slag is in a granulated form, roughly 500 microns in size, and has the following composition Fe 2 0 3 30.4 MgO 2.43 SiO 2 18.8 A1 2 0 3 3.74 CaO 22.1 MnO 0.20 ZnO 14.4 TiO 2 0.17

K

2 0 1.37 Due to the large consumption of acid by the slag material, the acid was sulphuric acid and was produced from sulphur dioxide and air. The sulphur dioxide was collected from the off gas from a copper smelting or lead sinter plant.

The ratio of sulphur dioxide to air determines the extent of oxidation of ferrous to ferric. In an aqueous acidic solution (the leach solution) sulphur dioxide is much more soluble than oxygen. In the absence of water, sulphur dioxide and oxygen will oxidise the ferrous iron to ferric iron. In the absence of oxygen and in the presence of water, sulphur dioxide will react with a ferric iron to form the ferrous iron and sulphuric acid.

Therefore, to balance the two reactions, a 0 preferred ratio is about 16% sulphur dioxide in air. A range of between 6% 20% sulphur dioxide can also be used but a preferred range is between 15% 16%.

As the solubility of oxygen in aqueous acid decreases with increasing temperature, it is also preferred that the temperature in the leaching step is maintained at between 40 0 C 70 0 C and typically at about 0

C.

To produce sulphuric acid in solution using sulphur dioxide and air, it is advantageous to add iron (typically as ferrous iron), to the starting leach solution. The iron can be introduced using FeSO.7H 2 0O and other iron Sources such as scrap iron and iron filings can also be used.

The amount of iron in the leach solution can range from 0.lg to up to 20g per litre or more. However, adequate production of sulphuric acid is obtained with a ferrous iron content of 2g 5g per litre. In this range, sulphuric acid is produced at an acceptable rate but without adding large additional inputs of iron which will require removal at later stages.

To provide an adequate supply of sulphur dioxide and air for leaching and acid production, and to keep a good efficiency of conversion of sulphur dioxide to sulphuric acid, a gas flow rate of between 0.3vvm (aeration volume per effective volume per minute) is acceptable with a preferred gas flow rate being lvvm. A dual impeller gas flow device can be used having a rushton turbine for good gas dispersion, and an axial impeller for o mixing.

The percentage of solid used in the leach slurry can vary between from 4%w/w to 25%w/w or more. Generally, S: a lower solids concentration will give a higher zinc *o S.extraction, however large volumes of slurry need to be used to process a reasonable tonnage of slag. Thus, a solids concentration of about 20%w/w provides an 30 acceptable extraction rate in a similar time frame and eeoc allows processing of greater amounts of slag. For instance, a solid content of 5% gives about 91% zinc extraction while a solids content of about 25% gives a 67% zinc extraction using equal process parameters.

The pH of the leach solution is important in terms of zinc extraction and impurity removal. Typically, the more acidic the solution, the greater the percentage of zinc extraction, and the lower the amount of formation -of colloidal silica. Thus, a pH of approximately 0.5 gives good extractions of zinc (about 90%) without a large amount of colloidal silica being formed.

However, the more acidic the leach solution is, the greater will be the amount of neutralisation material which is required to determine the optimum finishing pH.

The temperature of the leach solution can vary.

A temperature of between 40 0 C 70 0 C and preferably between 40 0 C 60 0 C can be used as higher temperatures reduces the solubility of oxygen in the leach solution.

The leach time can also vary and a typical leach time can be between 1 hour to 24 hours with a preferred time being between 3 5 hours.

Subsequent to leaching of the slag, the aqueous leach solution is removed and passed into a precipitation stage. Due to the mineralogy of the blast furnace slag, the zinc is mostly in the form of zinc ferrite and zinc silicate as well as some other oxides. The acidic leach solution also dissolves the iron in the slag and silica 20 as well as other impurities such as cadmium.

eeoce Iron must be removed prior to a solvent extraction as it will poison some organic reagents.

Silica must be removed as it forms a crud layer in the solvent extraction step.

25 The silica in solution has a tendency to form colloidal silica and silica gel which makes separation very difficult, and reduces zinc recovery levels.

However, it has now been found that if the iron in solution is precipitated as goethite, the removal of silica in solution is virtually total. It appears that simultaneous precipitation occurs of iron as goethite and colloidal or dissolved silica as monomeric anhydrous silica.

To ensure that iron in the leach solution precipitates as goethite and not as ferric hydroxide, the iron level in solution should be kept to below 2g per litre and- preferably below ig per litre. This can be achieved using pumping and automatic dosing equipment.

In the precipitation stage, the leach solution should be maintained at a temperature required to precipitate goethite instead of ferric hydroxide, the latter being more difficult to filter and having slower settling properties. The preferred leach solution temperature is greater than 75C and can be between 0 C 90 0 C and preferably about 80 0 C 85 0 C. At this temperature, it is thought that colloidal silica in solution precipitates in an anhydrous form.

To precipitate goethite from solution, it is desirable that the pH of the leach solution in the precipitation stage is above 3 and preferably above 3.8 to reduce goethite re-dissolving into the solution. The pH may be adjusted to the desired level by introducing a neutralising agent such as limestone.

As the majority of iron in the leach solution is present as ferrous iron, and as goethite is only formed from ferric iron, it is necessary to ensure that sufficient ferric iron is present. This can be achieved by oxidation of the ferrous iron and air can be used as ru the oxidant.

0 An air flow rate of lvvm into the leach

S..

solution was chosen to give adequate oxidation without excessive evaporation of the solution.

The residence time in the precipitation stage can vary and a typical residence time is between 5 hours with a preferred residence time being approximately hours. For larger scale operations, more reactors can be used in series increasing efficiency of goethite formation while keeping the overall residence time at about 10 hours. The residence time is also sufficient to e 4 6 convert ferrous iron to ferric iron in the precipitation stage, as it appears that the oxidation stage is the rate determining factor for iron removal.

A number of precipitation stages can be used in series. If more than one precipitation stage is used, these stages are usually at progressively higher pH values as the limestone is consumed, with the pH in the first vessel being 3.8.

The neutralising agent used in the precipitation stage can vary. Limestone for instance, in the form of a limestone slurry, can be used with an automatic controller doser. If limestone is used, gypsum will be formed as a by-product which will require disposal. Another neutralising agent which may be used is ZnO fume.

Using the goethite/silica precipitation step, about 92% of iron can be removed from solution with removal of up to 99.9% of silica (for copper recovery, only sufficient iron needs to be precipitated to remove all the silica) Figure 1 gives a schematic flow chart of a 15 preferred method of recovering zinc from a lead blast furnace slag. Initially, sulphur dioxide gas 10 from a

S..

copper or lead smelting process is cooled and cleaned 11 prior to acid production. The acid 12 and storage 13 is added to the starting slurry in tank 14. Slag at a 10% 20% slurry is leached in tank 12 with agitation and at a temperature of about 45 0 C, the leaching being exothermic.

The cleared off gas is also introduced into tank 14, together with ferrous ions. The leaching step can be batch or continuous and for large through rates, several 25 leach vessels having a retention time of up to 10 hours at 45 0 C can be used. The leach slurry feeds into the first of 4 precipitation stages 15 18 with a retention *time of 10 hours and a temperature of about 80 0

C.

Limestone slurry 19 from limestone mill 20 and zinc oxide 21 are added to the first precipitation stage 13 to neutralise the slurry to pH 4.5. It is important to note that no filtering is required between leaching tank 12 and precipitation stage 13 and thus the fact that the leach solution has silica in the form of a hydrated colloid is of little consequence. At stage 15, iron precipitates as goethite which filters easily and almost total co-precipitation of silica occurs during the iron removal. Gypsum also precipitates and the goethite/silica/gypsum precipitate is filtered off 22 and goes to a tailings dam 23. The filtrate, containing zinc sulphate and essentially free of silica and having acceptable levels of iron, goes to a temporary storage vessel 24 prior to solvent extraction by conventional techniques and electrowinning 25 by conventional techniques. The raffinate 26 is returned to acid production.

Copper may be recovered from copper smelter converter slag by the same method as described above except that during the precipitation step, the pH should not exceed 3.0 and preferably be in the range of 2.0 It should be appreciated that various other 15 changes and modifications may be made to the embodiment described without departing from the spirit or scope of the invention.

Claims

2. The method of claim 1 wherein the amount of ferric ion is less than about 2g per litre.
3. The method of claim 1 or claim 2 wherein the ferric ion is formed by introducing an oxygen containing gas into the solution to oxidise ferrous ion to ferric ion. 15 4. The method of claim 3 wherein the gas is air.
5. The method of any one of the preceding claims wherein the temperature of the solution during precipitation is about 750 to about 900
6. The method of claim 5 wherein the temperature is about 800 to about 850.
7. The method of any one of the preceding claims wherein the metal is zinc and the pH of the solution during precipitation is adjusted to be above about 3.
8. The method of claim 7 wherein the pH is above 25 about 3.8.
9. The method of claim 7 or claim 8 wherein substantially all the iron is precipitated. The method of any one of claims 1 to 6 wherein the metal is copper and the pH of the solution during precipitation is between about 2 to about 3.
11. The method of any one of claims 7 to 10 wherein the pH is adjusted by addition of a neutralising agent.
12. The method of claim 11 wherein the neutralising agent comprises limestone, zinc oxide or a mixture thereof.
13. A method for recovering a metal from a material comprising the metal and silica, said method comprising leaching the material with a leach solution comprising an aqueous acid to dissolve the metal and the silica, removing the silica by the method of any one of the preceding claims and separating the metal from the leaching solution.
14. The method of claim 13 wherein the material is a lead blast furnace slag comprising zinc or a copper containing slag. The method of claim 13 or claim 14 wherein the material further comprises iron.
16. The method of any one of claims 13 to wherein iron is added to the leach solution.
17. The method of any one of claims 13 to 16 wherein the pH of the leaching solution during leaching is about 15 18. The method of any one of claims 13 to 17 wherein the acid is sulphuric acid.
19. The method of claim 18 wherein a mixture of sulphur dioxide and oxygen and a ferrous catalyst are introduced into the leach solution such that the sulphur dioxide and oxygen react to produce sulphuric acid in the leach solution. :20. The method of claim 19 wherein sulphuric acid is added to the leach solution in addition to producing sulphuric acid in the leach solution.
21. The method of claim 20 wherein the added sulphuric acid is produced by reacting a mixture of sulphur dioxide and oxygen with a ferrous catalyst.
22. The method of any one of claims 19 to 21 wherein the mixture of sulphur dioxide and oxygen is a sulphur dioxide/air mixture comprising about 3% to about sulphur dioxide.
23. The method of claim 22 wherein the mixture comprises 16% sulphur dioxide.
24. The method of any one of claims 19 to 23 wherein the sulphur dioxide mixture comprises an off gas from a copper smelter, a lead smelter or from burning sulphur. The method of any one of claims 19 to 24 14 wherein the mixture is introduced into the leach solution at a flow rate of between 0.3vvm to
26. The method of any one of claims 13 to wherein the leach solution has a solids concentration of between about 4 and
27. The method of any one of claims 13 to wherein the metal is separated by solvent extraction followed by electrowinning. 22. A method of recovering a metal from a material comprising the metal and silica, substantially as hereinbefore described with reference to the Figure. DATED this 14th day of July 1997 M.I.M. HOLDINGS LIMITED 15 009 814 019) By their Patent Attorneys CULLEN CO. 0. 0 *e.o