A process is disclosed for the treatment of industrial residues/wastes for the recovery of metals. Such residues/wastes are in the form of difficult to process solids or sludges. The process can recover metals such as lead and zinc, as well as other metals such as copper, nickel, cobalt, silver, mercury, etc.

Background Art

Heavy metals are extensively used in modern society for the production of commercial and domestic goods (e.g. cars, electronics, building materials, batteries, etc). Despite ongoing improvements to processes used in manufacturing these goods, heavy metals still pass into waste and residue streams from these processes. Increasingly, these wastes must be dealt with in an environmentally and sustainable fashion. For example, the metals can be precipitated from waste solutions (although sometimes the waste is already a solid), with the resulting cake immobilised in landfill sites.

The above references to the background art do not constitute an admission that the art forms a part of the common general knowledge of a person of ordinary skill in the art. The above references are also not intended to limit the application of the process disclosed herein.

Summary of the Disclosure

In a first aspect there is disclosed a process for recovering one or more metals from an industrial residue/waste. The process comprises the steps of:

(a) adding the residue/waste as a sludge or solid to a leaching stage in which it is leached with an aqueous halide solution that has an oxidation potential and acidity sufficient to render soluble in the solution one or metals present in the sludge or solid;

(b) passing the solution from (a) to a precipitation stage in which the one or more metals are precipitated as a metal sulfide; (c) passing the solution from (b) to a regeneration stage in which the sufficient oxidation potential and acidity is regenerated in the aqueous halide solution for recycle to the leaching stage (a).

Such a process can be used to recover, as a metal sulfide, otherwise difficult-to-recover metals from industrial sludges and solid wastes/residues, many of which metals may have value. The industrial residues/wastes that can be treated using the process include, but are not limited to, metal oxides, hydroxides, sulfides, and carbonates (hereinafter "metalliferous materials"). The residues/wastes may come from a variety of industries such as automotive, marine, mining and metals processing, steel, construction, electroplating, electronics and analytical services industries.

It should be noted that where the metalliferous material itself already comprises a sulfide, the process is nevertheless also able to separate out that sulfide as one or more specific target metal sulfides in the precipitation stage (b) (i.e. for subsequent recovery of the target metal). For example, where the metalliferous material contains iron the process can separate a metal sulfide from the iron.

The metal(s) from the metal sulfides produced in the process can be recovered in a known manner (e.g. by smelting or roasting), whereby a sulfuric acid by-product of the smelting or roasting process can be returned for use in acid regeneration stage (c) of the process.

The process is also able to be operated in an efficient closed loop configuration whereby the acidic aqueous halide solution that leaches the residue/waste is able to be subjected to sulfidisation in the precipitation stage (b) and can then be regenerated in stage (c) for recycle to and reuse in the leaching stage (a). Furthermore, each of the leaching, precipitation and regeneration stages can be provided as circuits which, in the process, are able to be integrated.

The process is also able to bring together reagent production, regeneration and consumption. This can considerably improve process economics and may also result in no hazardous waste being produced (e.g. a waste that requires any further treatment prior to subsequent reuse or disposal).

In one embodiment of the process, in precipitation step (b), the one or more metals can be sulfidised by the addition to the solution of a different metal sulfide (i.e. comprising a metal that is different to the target metal(s) leached in step (a)). For example, the different metal can be calcium, which has been found to be the most suitable alkaline earth metal in this regard (though conceptually another alkaline earth metal may be used).

In the case where the different metal is calcium, calcium sulfide can be added to the precipitation stage (b) to precipitate the one or more metals as a metal sulfide. The calcium sulfide can be sourced separately; however, as explained below, the process can be operated whereby the calcium sulfide is produced from a calcium sulfate precipitate produced in the acid regeneration stage. In either case, the metal sulfides produced in precipitation stage (b) can then be separated and recovered from the solution (e.g. through the use of a thickening stage and/or a filtration stage), whereas the calcium can pass with the solution to the acid regeneration stage (c).

Usually a generally "clear" (low or particulate-free) solution from the leaching stage (a) is fed to the precipitation stage (b). In stage (b) the solution is sulfidised to recover the now dissolved target metal(s). The calcium sulfide added to precipitation stage (b) can also react with residual acid in solution to form hydrogen sulfide (H ₂ S), which also reacts with metal-halide complexes in solution to precipitate metal sulfides. The solution pH in stage (b) is typically less than 7 and may be in the range of 2-4.5, being a range that favours production of a stable target metal sulfide precipitate. The solution temperature in stage (b) can be between 25-115°C, optimally around 80°C. The solution residence time can range from 0.5-24 hours at atmospheric pressure, and may be around 7 hours.

In one embodiment of the process, in acid regeneration step (c), the acidic aqueous halide solution is regenerated by the addition to the solution of sulfuric acid. For example, where the solution halide is chloride and/or bromide then hydrochloric acid and/or hydrobromic acid is regenerated. In the case where the different metal is calcium, the addition of sulfuric acid also results in the calcium in solution being precipitated as calcium sulfate.

In this embodiment of the process, the regeneration stage (c) can be controlled to favour the production of anhydrous calcium sulfate. In this regard, the exothermic reaction that results from the addition of sulfuric acid can be maximised, with increased solution temperatures (for example, in the range of 50-115°C) favouring anhydrous calcium sulfate (anhydrite) formation over the bassanite or hemihydrate (CaS0 ₄ .½H ₂ 0) and gypsum (CaS0 ₄ .2H ₂ 0) forms. These increased solution temperatures can also be "recycled" to favour the leaching stage (a).

It is further noted that the particle shape and size of calcium sulfate (anhydrite) produced in the regeneration stage (c) can be further controlled (e.g. through the use of high pressure and temperature vessels, such as an autoclave). Thus, a high value calcium sulfate may be produced (e.g. a so-called calcium sulfate "whisker").

In this embodiment of the process, prior to recycling the acidic aqueous halide solution to the leaching stage (a), the calcium sulfate precipitate can be separated from the solution (e.g. through the use of a thickening stage and (or only) a filtration stage).

In an embodiment the process may then comprise an additional step (d) of reducing the separated calcium sulfate to calcium sulfide. Advantageously, this calcium sulfide may then be recycled for use as the different metal sulfide in the precipitation stage (b). In this way the process brings together reagent production, regeneration and consumption.

In this embodiment the calcium sulfate can be reduced to calcium sulfide for recycle to the precipitation stage (b) by contacting the calcium sulfate with a reductant that is capable of reducing calcium sulfate to calcium sulfide. For example, the reductant can be selected from the group comprising a source of carbon or hydrogen. The source of carbon may take the form of a naturally occurring gas such as "natural gas" or methane, or a low-ash carbon such as an organic or inorganic ash, carbon from an industrial/natural waste or by-product including from biomass, a fuel oil etc.

In additional step (d) the calcium sulfide can be produced in a furnace, where the calcium sulfate is converted to calcium sulfide in a reducing atmosphere at a temperature ranging between 600-1200°C.

In one embodiment of the process the leaching stage (a) can be controlled to favour the precipitation of iron (when present) in the residue/waste. This iron, together with a leach residue, can then be separated from the solution prior to passing it to the precipitation stage (b). In this regard, the oxidation potential (Eh) in the leaching stage (a) can be optimised to such that Fe is in its ferric Fe ³⁺ form to favour Fe precipitation.

For example, where the leaching stage (a) comprises multiple leach reactor stages the first and second stages can be controlled to have a high Eh in the range of 450-650 mV (ref Ag/AgCl) to promote the oxidation of Fe ²⁺ to Fe ³⁺ . Then, in a subsequent reactor stage Fe ³⁺ can be precipitated as Fe ₂ 0 ₃ (hematite) by raising the solution pH (e.g. above pH 1.2). The precipitated iron passes out with the leach residue. This "purified" and stable residue may be recycled to an industrial process or disposed of. Depending on the residue/waste to be treated, the temperature for the leaching stage (a) can range from 20-1 15°C, the residence time can range from 0.5-24 hours, and at least the initial reactor stages can have a pH of around 1 or less. To achieve higher leach solution temperatures, the process can make use of the exotherm produced in the acid regeneration stage (c), whereby the solution recycled to leaching stage (a) may be at temperatures elevated to greater than 60°C. Also, when the metalliferous material is an oxide, this is able to be leached faster (e.g. around 1 hr or less) than a sulfide (which can require residence times of a number of hours).

Some metalliferous materials may comprise non-sulfide forming metals such as gold. In such cases, in addition to the precipitation stage (b), the process may comprise an additional metal recovery stage (e) for recovering the non-sulfidising metals from the solution. For example, the metal recovery stage (e) can comprise one or more of an ion-exchange resin stage and/or a solvent extraction stage. These stages may also take the form of circuits that can be integrated with a precipitation circuit (b). For more difficult-to-recover metals such as indium, the ion-exchange resin can be a chelating resin (a subgroup of ion-exchange resins).

In one embodiment of the process each of the leaching stage (a), precipitation stage (b) and regeneration stage (c) comprises multiple reactor stages operated in a co-current

configuration. Employing multiple reaction stages allows for better control of each of the stages (a), (b) and (c), generally resulting in improved yields, and better targeting of specific impurities or to-be-recovered metals. A co-current configuration allows for better integration of the flow circuits between the leaching stage (a), precipitation stage (b) and regeneration stage (c), with minimal solid/liquid separation equipment required.

However, in some applications of the process a counter-current configuration may be adopted for the leach circuit. This configuration may be required where the specific

residues/wastes are complex. The counter-current configuration can achieve selective leaching of component minerals and chemical compounds, and different leach conditions can be more readily applied than in a co-current configuration of multiple reactors. This can be most evident in cases where reduced acid consumption or reduction-oxidation potential control is desired.

In one embodiment of the process each of the stages (a), (b) and (c) can comprise a final separation stage for removal from the solution of resultant solids, before the solution is passed on to the next stage. In each of stages (a), (b) and (c) the final separation stage may comprise a thickening stage and/or a filtration stage for solids removal. However, in some applications (e.g. in the acid regeneration stage (c)) a thickening stage may not be required.

The separation stage can take the form of a solid-liquid separation circuit. In this regard, the slurry may be first sent to a thickener; with the resulting underflow slurry then forwarded to a filter for recovery of solids. The filtered solids can be washed and, in the case of:

(a) the leaching stage - the leach residue can be disposed of or recycled to industry;

(b) the precipitation stage - the metal sulfides (which constitute the "product" of the process) can be recovered in a known manner (e.g. by smelting or roasting to recover the target metal(s));

(c) the regeneration stage - the metal sulfate precipitate (e.g. calcium sulfate) is produced as a saleable product and/or is used in additional step (d) (i.e. to be reduced to calcium sulfide).

In the case of additional step (d) the calcium sulfide product may be sold and/or recycled as a reagent to precipitation stage (b).

Usually just a small proportion of the calcium sulfate and calcium sulfide are sold, with the bulk (or entirety) reused in the process itself. However, by adjusting the conditions in the stages (b) and (c), value-added forms of both calcium sulfate and calcium sulfide can be produced.

The thickening stage can make use of high rate thickeners, low rate thickeners, clarifiers and similar devices for solid-liquid separation. The filtration stage can make use of pressure filters, pan filters, belt filters, press filters, centrifuge filters and similar devices for solid-liquid separation.

In one embodiment of the process each of the stages (a), (b) and (c) can comprise one or more recycle streams to allow for control of solids residence time in the solution as well as to improve yield/recovery. The (or each) recycle stream can be from a given reactor stage to a previous reactor stage; a so-called "internal" recycle (e.g. the slurry from one reactor is recycled back to a previous reactor). Alternatively or additionally, the (or each) recycle stream can be from the separation stage (e.g. from the thickener slurry underflow) to a given reactor stage; a so- called "external" recycle.

In one embodiment of the process the acidic aqueous halide solution comprises chloride, or a mixture of chloride and bromide. The acidic aqueous halide solution usually comprises a soluble metal halide solution. The metal halide concentration can be 1-8 moles per litre (e.g. comprising one or a mixture of NaCl, NaBr, CaCl ₂ , and CaBr ₂ ). However, the use of other halides is possible.

In one embodiment of the process the acidic aqueous halide solution may further comprise a multi-valent species that is able to provide a reduction-oxidation couple to facilitate oxygen exchange into solution. For example, a multivalent metal (such as copper or iron) can be introduced by the residue/waste, or can be deliberately added, to provide a species that exchanges oxygen into the system. The oxygen can be sparged into the solution as air.

The particular leaching solution employed in this process enables a much wider range of residues/wastes to be treated via a hydrometallurgical route compared to the known lixiviants sulfuric acid, nitric acid, ammonia, or caustic-solution leaches.

BRIEF DESCRIPTION OF THE DRAWINGS

Notwithstanding any other forms which may fall within the scope of the process as defined in the Summary, specific embodiments will now be described, by way of example only, with reference to the accompanying drawings in which:

Figure 1 shows a block diagram for an embodiment of the process comprising a number of circuits that are integrated to process metal residues/wastes into, inter alia, a metal sulfide product;

Figure 2 shows a flowsheet for an embodiment of the leaching circuit of the process of Figure 1;

Figure 3 shows a flowsheet for an embodiment of the precipitation (sulfidisation) circuit of the process of Figure 1;

Figure 4 shows a flowsheet for an embodiment of the acid regeneration circuit of the process of Figure 1;

Figure 5 is a graph showing the particle size distribution of CaS0 ₄ produced using the acid regeneration methodology of Example 1; and

Figure 6 is a plot showing the X-ray diffraction analysis of CaS0 ₄ produced using the acid regeneration methodology of Example 1. Detailed Description of the Specific Embodiments

Figure 1 shows a process block diagram for recovering one or more metals from an industrial residue/waste. The flowsheet of Figure 1 comprises three main integrated circuits: a leach circuit 100, followed by a metal sulfide precipitation (or sulfidisation) circuit 200, and then an acid generation (or regeneration) circuit 300. Optionally (and advantageously) the process can comprise a sulfate-to-sulfide reduction circuit 400. As a further option, the process can comprise a metal recovery circuit (not shown) for the recovery of non-sulfide forming metals that have been leached into solution. The metal recovery circuit can comprise one or more of an ion- exchange resin circuit and/or a solvent extraction circuit. The metal recovery circuit can be integrated into the precipitation (sulfidisation) circuit 200.

The flowsheet depicts a closed-loop process whereby the leaching, precipitation and regeneration stages (circuits) are integrated to bring together reagent production, regeneration and consumption. This improves process flow characteristics and process economics including the deployment of capital.

More particularly in the process, the residue/waste is fed as a sludge or solid stream 101 to the leach circuit 100 where the metal oxides, hydroxides, sulfides, and/or carbonates (hereafter "metalliferous material") are solubilised using a hydrohalic acid such as hydrochloric and/or hydrobromic acid. In Reaction 1 (below) the use of hydrochloric acid (being a typical acid employed in the process) is represented. The resultant leached solids are separated as a leach residue stream 103 at separation stage 110. The separated liquid solution (i.e. leachate) is passed as stream 104 to the metal sulfide precipitation circuit 200.

Dissolved metals in stream 104 are precipitated (or otherwise separated) in the sulfidisation circuit 200. This circuit makes use of a metal sulfide as the sulfidisation reagent. In Reaction 2 (below) the use of calcium sulfide (being a typical metal sulfide employed in the process) is represented. In this regard, the metal sulfide may be separately purchased as a reagent and fed in to circuit 200 as stream 403. However, in accordance with the integrated configuration of the present process, optimally the stream 403 is a recycle of the metal sulfide produced in the sulfate-to-sulfide reduction circuit 400. Usually the stream 403 comprises calcium sulfide.

The precipitated target metal sulfides (and other separated metals) are separated as a product stream 202 at separation stage 210. Target metals in the precipitated sulfides include lead, zinc, copper, nickel, cobalt, silver, mercury etc. Target metals from the ion-exchange resin and/or a solvent extraction circuits include gold. The separated solution (liquor) is passed as stream 203 to the acid regeneration circuit 300. The metals of the metal sulfides (and in other formats) are recovered in a known manner, such as being on-sold for smelting or roasting. When the metals are recovered by smelting or roasting usually such processes produce a sulfuric acid by-product. This can be returned for use in the acid regeneration circuit 300 as stream 301.

In acid regeneration circuit 300 the hydrohalic acid (usually hydrochloric and/or hydrobromic acid) is regenerated by reacting the overflow liquor 203 with concentrated sulfuric acid fed in as stream 301. In Reaction 3 (below) this is shown as the regeneration of hydrochloric acid (being a typical acid employed in the process) as well as the precipitation of calcium sulfate (i.e. when calcium sulfide is added at precipitation stage 200). The calcium sulfate produced is separated as stream 303 at separation stage 310. The separated regenerated acidic solution (i.e. regenerated lixiviant) is recycled as stream 306 to the leach circuit 100.

Whilst a portion of the calcium sulfate produced in acid regeneration stage 300 can be sold (i.e. as stream 305), usually a bulk (if not all) of the calcium sulfate is fed as stream 304 to the sulfate-to-sulfide reduction circuit 400 for the production of calcium sulfide reagent 403.

Acid regeneration in circuit 300 is optimally performed at higher temperatures, making maximum use of the exotherm of Reaction 3. This can ensure production of a high purity, highly crystalline anhydrous calcium sulfate (i.e. anhydrite). Optionally, acid regeneration can be performed in a special high pressure and temperature vessel (e.g. an autoclave). This can allow the particle shape and size of calcium sulfate produced to be further and better controlled, in turn producing an even higher value calcium sulfate (e.g. a so-called calcium sulfate "whisker").

In circuit 400 the source of carbon is fed into a furnace as stream 401, along with the metal sulfate stream 304. This is shown in Reaction 4 (below), where the reductant is indicated as a source of carbon C, although another reductant (e.g. a reducing gas such as hydrogen) can be employed. The carbon source may be a naturally occurring gas such as natural gas, methane, a low-ash carbon such as an organic or inorganic ash, carbon from an industrial/natural waste or by-product including from biomass, a fuel oil etc.

In the furnace the calcium sulfate is converted to calcium sulfide in a reducing atmosphere at a temperature ranging between 600-1200°C. When the reductant is a carbon source a C0 ₂ by-product stream 402 is produced along with the metal sulfide streams 403 and 404. When the reductant is a gas such as hydrogen an H ₂ 0 by-product stream 402 can be produced along with the metal sulfide streams 403 and 404.

The chemistry Of the four circuits can be summarised as follows: Leach MO _(j) + HCl _(/) -» MCl _2(/) + H ₂ 0 _(/) Reaction 1

Sulfidisation MC\ ₂ (D + CaS _(j) - MS _W + CaC½ Reaction 2

Acid Regeneration H ₂ S0 _4(¾¾ > +CaCl ₂ .nH ₂ 0^ _aft ) - CaS0 ₄ .nH ₂ 0^ + 2HC1^ ₉ Reaction 3

Sulfate Reduction CaS0 _i) + 4C ₍₅₎ - CaS^ ₎ + 4CO _(G) Reaction 4

Leach Circuit (Fig 2)

Referring now to Figure 2, the leach circuit 100 is shown as comprising five cascading reactors (Reactor 1 to Reactor 5), followed by a separation stage (110) comprising a thickener, a belt filter, and a surge tank to collect the overflow process liquor. The number of reactors in leach circuit 100 may vary down to as simple as one reactor, depending on the waste/residue to be leached. Other circuit configurations are not precluded, and other configurations employed will depend on the chemical complexity of the metalliferous material.

In leach circuit 100 the metalliferous material (stream 101) is mixed with hydrochloric acid (stream 306) that has been recycled from acid generation circuit 300. The mix in Reactor 1 is controlled to obtain a slurry density ranging from 1-30 % w/w, usually around 10-15 % w/w. In Reactors 1 through to 5 the oxidation-reduction potential Eh is maintained sufficiently high to dissolve the metal content and maintain the solubility of the metal cations. An Eh of >200 mV (versus Ag/AgCl) and preferably in the range of 450-650 mV is maintained. In order to increase solution Eh in the leaching step, additional oxidants are added as necessary such as oxygen, air, chlorine gas (all may be sparged in), or a halide complex such as BrCl ₂ ^" is generated (in an electrolytic cell) and added.

In selected Reactors the Eh is maintained in the range of 500-650 mV to promote the oxidation of iron (when present) in the metalliferous material from Fe ²⁺ to Fe ³⁺ . In the subsequent Reactors the solution pH is increased (i.e. greater than around 1.2) whereby the Fe ³⁺ may be readily precipitated as Fe ₂ 0 ₃ (hematite) to ultimately pass out with the leach residue stream 103.

In leach circuit 100 leaching is carried out at a temperature in the range of 20-115°C, more preferably at 60-80°C, and for a residence time of 0.5-24 hours under atmospheric pressure, depending on the nature of the metalliferous material. For example, metalliferous oxides oxidise and leach typically in less than 1 hr, whereas metalliferous sulfides oxidise and leach typically over many hours, and can require considerable recycle.

To control solids residence time and improve yield/recovery of target metal(s) from the leaching stage, the circuit 100 comprises a number of recycle streams. In Figure 2 there are two so-called "internal" recycle streams depicted where slurry is recycled from Reactor 3 to Reactor 2, and from Reactor 5 to Reactor 4. In addition, a so-called "external" recycle stream of slurry underflow from the thickener to each of Reactor 4 and Reactor 2 is shown.

The resultant slurry overflow from Reactor 5 (stream 102) is fed into the solid-liquid separation system 110 which comprises a thickener and a filter. In the thickener the solids are increased to a density ranging from 20-80 % w/w, usually around 30-40 % w/w. Appropriate flocculants and coagulants 106 may be added to the slurry fed to the thickener to improve the efficiency of the solid-liquid separation steps.

Thickener overflow liquor, comprising <1% suspended solids, is collected in a stock/surge tank prior to being forwarded as stream 104 to target metal recovery (circuit 200). The underflow slurry is sent to a belt filter where four counter-current stages of washing with fresh water are performed. The primary filtrate from the belt filter is recombined with the thickener overflow in the stock tank to produce stream 104, while the leach residue (stream 103) is discharged. Depending on the content of the original metalliferous material the leach residue stream 103 may be recycled to an industrial process, or disposed of as a relatively non-toxic residue to landfill.

Sulfidisation (Precipitation) Circuit (Fig 3)

Referring now to Figure 3, the metal sulfide precipitation (sulfidisation) circuit 200 is shown as comprising three cascading reactors (Reactor 1 to Reactor 3), followed by a separation stage (210) comprising a thickener, a filter, and a surge tank to collect the overflow process liquor. Again, the number of reactors and the circuit configuration in sulfidisation circuit 200 may vary depending on the chemistry of the leachate, which in turn may depend on the chemical complexity of the metalliferous material.

Should there be non-sulfide forming metals in the leachate, ion-exchange or solvent extraction circuits maybe be included on either stream 104 or 203.

In the sulfidisation circuit 200 the clear liquor advancing from the leach circuit (stream

104) is fed to Reactor 1 where it is contacted with an internal recycle from Reactor 3. A calcium sulfide stream 403 is introduced into Reactor 2 to promote a reaction with one or more of the target metals in solution according to Reaction 2 (above). The calcium sulfide added to Reactor 2 is also able to react with residual acid in the liquor (i.e. from the leach circuit) to form hydrogen sulfide (H ₂ S). Hydrogen sulfide also reacts with metal-halide complexes in the liquor to precipitate the metal sulfides.

The solution pH in circuit 200 is less than 7 and is typically controlled to be in the optimal sulfide precipitation range of 2 - 4.5 by adjusting calcium sulfide addition, and/or by employing acid or neutralising agents. For example, where residual acid is too high in the liquor from the leach circuit (stream 104) it can be reduced by the addition of CaC0 ₃ . The pH range of 2 - 4.5 favours the production of a stable target metal sulfide precipitate. The solution temperature in circuit can be between 25-115°C, preferably in the range 60-80°C, and optimally around 80°C. The solution residence time can range from 0.5-24 hours at atmospheric pressure, and may be around 7 hours.

In circuit 200 a number of recycles are employed; the process liquor is fed to Reactor 1, slurry is internally recycled from Reactor 3 back to Reactor 1, CaS reagent is added to Reactor 2 as stream 403, and the Reactor 3 is employed to increase residence time and to enable pH adjustment. This system improves particle settling and filtration characteristics by preferentially forming more crystalline and regular sized target metal sulfide particles. It also helps increase reagent utilisation by allowing better control of the pH and Eh conditions.

The slurry exiting the reactors is then sent to a solid-liquid separation step 210. Typically a thickener is employed followed by solid filtration using a belt or pressure filter, in a similar manner to the leach circuit 100 (described above). The resultant liquor 203 to be sent to acid regeneration now comprises dissolved calcium, is depleted of target metals and is essentially depleted of particulates.

Acid Regeneration Circuit (Fig 4) Referring now to Figure 4, the acid regeneration circuit 300 receives process liquor from the sulfidisation circuit 200 containing calcium chloride (stream 203) and reacts this liquor with concentrated sulfuric acid (H ₂ S0 ₄ ) (stream 301). This chemical reaction is shown in Reaction 3 (above), and results in the precipitation of calcium sulfate as a feed to the reduction circuit 400 (stream 304) and a by-product (stream 305).

The acid regeneration circuit 300 comprises two cascading reactors (Reactor 1 and Reactor 2), followed by a separation stage (210) comprising a filter, and a surge tank to collect the overflow regenerated liquor. Again, the number of reactors and the circuit configuration in circuit 300 may vary depending on the chemical complexity of the metalliferous material.

In Reactor 1 the pH is typically controlled to be between 0 and 1 and the reaction temperature is kept between 50°C and a maximum of 115°C (i.e. the approximate boiling point of the liquor). Under atmospheric conditions, the temperature of the circuit is controlled by the highly exothermic reaction that occurs when concentrated sulfuric acid (~ 766 kJ/kg for 98 % H ₂ S0 _4(aq) ) is diluted. The circuit takes advantage of these conditions to produce the preferred form of calcium sulfate as anhydrite (CaS0 ₄ ), which is highly crystalline and filterable when compared to gypsum (CaS0 ₄ .2H ₂ 0).

In circumstances where there is a demand for higher value forms of calcium sulfate, the conditions and equipment can be modified to produce niche products. These can include calcium sulfate whisker produced at 150-200°C in pressure vessels (e.g. an autoclave).

In circuit 300 it will be seen that there is an internal recycle from Reactor 2 to Reactor 1, a filter and a stock tank. The concentrated sulfuric acid is introduced into the Reactor 2, where it is mixed with the advancing slurry from the Reactor 1. In Reactor 1 the internal recycle stream delivers unreacted sulfuric acid to preheated feed solution (which is itself rich in calcium, post the metal sulfides recovery stage). Thus, Reactor 1 also produces hydrochloric acid and a calcium sulfate by-product.

Both Reactors 1 and 2 are designed to operate at elevated temperatures (90 - 110°C) using preheated feed solution coupled with the heat generated from sulfuric acid addition.

The advancing pulp from Reactor 2 is separated via a filter where the filtrate stream delivers the hydrochloric acid into an acid stock tank ready for recycle to the leach circuit 100 as stream 306. The calcium sulfate (stream 303) is recovered via the filter.

Reduction Circuit (Fig 1) The calcium sulfate (stream 303) from the acid regeneration circuit 300 is able to be split. A bulk (if not all) of the calcium sulfate (stream 304) is mixed with low-ash carbon (stream 401), and fed to a furnace 400 where the sulfate is converted to sulfide at 600 - 1200°C, and typically at 900-1000°C, in a reducing atmosphere. The resulting calcium sulfide (stream 403) is used as the reagent in the metal recovery circuit, although a portion (stream 404) can be sold.

Examples

Non-limiting Examples of various stages (circuits) of the process for recovering one or more metals from an industrial residue/waste will now be described.

Example 1: Demonstration Plant Acid Generation Trials

A demonstration plant was continuously operated for seven days where a total of 70m ³ of calcium rich process liquor (30.5 g/L) was reacted with 7.8 m ³ of 98 % H ₂ S0 ₄ ( _aq ) to generate 5.4 t of CaS0 _4(S) and 10.7 t of HC1 at a concentration of 129 g/L. The acid generation circuit comprised two reactors (with an internal recycle), a thickener and a stock tank. Both reactors were designed to operate at elevated temperatures (90 - 110°C) using preheated feed solution coupled with the heat generated from sulfuric acid addition (an extremely exothermic reaction).

The resulting calcium sulfate product was filtered and analysed, and was shown to contain a mixture of gypsum and anhydrite by X-ray diffraction, as per the following table. The particle size was p80 = 150 μπι.

Material Specifications

Calcium 26%

Sulfate 67%

Moisture 7%

As shown in Figure 5, the particle size distribution was calculated to be p80 = 150 μηι. Figure 6 shows the X-ray diffraction analysis of the resulting calcium sulfate product.

The following Table shows that filtration rates were high, with an average of ~ 800 kg/m ² /hr for three samples.

5

Example 2: EAF Dust Leach Trials

The leach step was evaluated by performing bench leach tests, pilot plant leach tests and continuous demonstration plant leach tests. The typical mineral composition for the EAF Dust (EAFD) treated is shown in the following Table:

Zincite ZnO 18 - 20

Magnetite Fe ₃ 0 ₄ 4 - 12

Calcite CaC0 ₃ 1 - 3

Hematite Fe ₂ 0 ₃ 0 - 4

Halite NaCl 2 - 3

Quartz Si0 ₂ 1 - 2

Sylvite KC1 <1

A demonstration plant was continuously operated for a period of 8 days to treat 30 tonnes of EAF Dust. The leach circuit consisted of five reactors operating at a temperature range of 95- 110°C, a pH range from 0.3 in the first reactor to 1.3 in the last reactor. Acid was forwarded to the leach circuit from the previous acid generation stage, as described in Example 1.

The results showed an acid consumption of only 430 kg of H^CVt EAFD. Metal extractions were between 77% and 96% as shown in the following Table.

Example 3: Demonstration Plant Sulfidisation Trials

Sulfidisation of contained zinc and lead in the leach process liquor from Example 2 was successfully achieved by adding calcium sulfide reagent. The circuit consisted of a feed tank, three cascade reactors followed by a thickener and a surge tank to collect the overflow process liquor. The demonstration plant was run for four continuous days with no major downtime. A total of 21 m ³ of leach process liquor, at an average concentration of 52 g/L zinc and 2 g/L lead, was fed to the circuit along with 2.2 t of calcium sulfide reagent (at 49% CaS), producing 1.6 t of dry zinc-lead sulfide product at an average grade of 34.6% zinc and 1.5% lead. The residence time for this circuit was 7.5 h and the reaction was conducted at a temperature of 80 °C and a reactor pH of 2 - 4.5 under atmospheric pressure.

Calcium sulfide utilisation was observed to be approximately 92.3%, with 5% being consumed by excess addition of hydrochloric acid into the circuit. To control the circuit pH, an acid consumption of 340 kg HCl/t CaS was noted, which was slightly elevated due to the presence of contaminant lime (CaO) within the CaS reagent. On a lime basis only, the acid consumption equated to 1.3 kg HCl/kg CaO, which was in accordance with chemical stoichiometry. This was equivalent to 93% of the total acid addition; hence the excess acid (of 7%) was consumed by the calcium sulfide side-reaction to form hydrogen sulfide gas, equivalent to 5,000 kg HCl/t CaS when determined stoichiometrically.

The Tables set forth below list the key performance indicators for the trials, and the zinc- lead sulfide mineralogy, respectively.

Compound Unit Value

ZnS % 50

PbS % 2.1

CuS % 0.2

AgaS ppm 37

CdS ppm 502

CaS0 ₄ .2H ₂ 0 % 6.1

A1 ₂ 0 ₃ % 5.1 Fe ₂ 0 ₃ % 1.6

C % 5.7

Si0 ₂ % 12.3

MnFe ₂ 0 ₄ % 3.8

MgS0 ₄ .7H ₂ 0 % 1.2

S (ele) % 1.9

NaCl % 0.2

KC1 % 0.4

Ti0 ₂ % 0.2

Total % 91

Example 4: Muffle Furnace Trials for Calcium Sulfate Reduction

The calcium sulfate prepared in Example 1 was converted into calcium sulfide using a laboratory sized muffle furnace. Two series of tests were conducted. The first were small scale tests where 10 g of calcium sulfate was reacted with carbon. Low ash carbon granules were physically mixed with the calcium sulfate, prior to heating the mixture in the furnace. As shown in the following Table, the product contained approximately 90% calcium sulfide. Larger tests were then conducted. The average results (next Table after the following) indicated that the total product grades were 72% calcium sulfide, 10% unreacted calcium sulfate, and 17% unreacted carbon. However, after size separation, the fine fraction graded 85% calcium sulfide and 6% carbon, while the oversize fraction graded 57% calcium sulfide and 30% carbon. This indicated that the conditions for reaction were highly dependant on the particle size of the carbon.

Table: Calcium Sulfate Reduction into Calcium Sulfide

Example 5 - Demonstration Plant Treatment Trial

A process for recovering lead from an industrial residue/waste was evaluated by treating 8300 kg of lead-containing carbonate sludge and waste water from a third-party manufacturing process. The solids component was observed to be -25% of the total waste (-2075 kg), and the composition of the hazardous waste material was as follows:

A solution of hydrochloric acid was prepared by reacting sulfuric acid and a calcium chloride solution. 5000 kg of the resulting acid solution, containing - 15% hydrochloric acid, was reacted with the waste sludge. The quantity of acid required for leaching was noted to be heavily dependant on the soda ash (Na ₂ C0 ₃ ) component, which in turn was noted to be linked to the waste generation process used by the third-party.

The pH of the leach solution was between 1 and 1.5. 850 kg of 50% calcium sulfide was then added to the solution. The Eh was maintained between 0 and -50 mV (vs Ag/AgCl). Three reactors were used for this sulfidisation step, and a sequence of recycle flows was employed to improve product crystallisation and reagent consumption. The total residence time was eight hours in the circuit.

On a dry basis, the weight of the recovered sulfide material was -2200 kg. The final product composition was: Element Wt%

Pb 38.7

Cu 3.7

Sn 8.1

Ni 5.9

Fe 3.2

Zn 0.6

Ca 8.9

Al 1.8

Na 0.7

S 10.5

Si0 ₂ 8.9

The Si0 ₂ and CaO components were noted to be residual minerals from the impurities in the calcium sulfide reagent. The use of higher grade calcium sulfide was observed to minimise these "contaminants" in the final metal sulfide product.

This example demonstrated the conversion of a hazardous lead carbonate waste into a valuable sulfide concentrate. The lead sulfide concentrate was able to be processed using conventional lead recovery techniques, including pyrometallurgical smelting. The process embodiments described herein was observed to provide a treatment technology for hazardous (heavy metal-containing) industrial wastes and residues that was economic, sustainable in the long-term, environmentally friendly and produced useful, valuable and non-toxic products and by-products.

Whilst a number of specific process embodiments have been described, it should be appreciated that the process may be embodied in other forms.

In the claims which follow, and in the preceding description, except where the context requires otherwise due to express language or necessary implication, the word "comprise" and variations such as "comprises" or "comprising" are used in an inclusive sense, i.e. to specify the presence of the stated features but not to preclude the presence or addition of further features in various embodiments of the process as disclosed herein.