RECOVERY OF ALUMINA FROM COAL FLY ASH

TECHNICAL FIELD

[001] This disclosure relates to a process for remediation of impounded coal ash, such as may be held in a storage pond.

BACKGROUND ART

[002] Coal ash is a by-product from the burning of coal in thermal power stations. There are millions of tonnes of coal ash that has been produced from the thermal power generation industry around the world. Notwithstanding that over the years some of this coal ash has been beneficially used, much of it is still stored in containment dams adjacent to the power station, regardless of whether the power station is still in operation.

[003] Most of this ash contains toxic heavy metals and metalloids. In Australia, it is estimated that 650 million tonnes of coal ash are in storage. In NSW alone, it is estimated that 216 million tonnes are in storage. Various studies have shown that toxic elements present in the coal ash have and are still leaching into the environment.

[004] Due to the presence of toxic heavy metals and metalloids within the coal ash, most coal ash is, at present, unable to be used unless special exemptions are granted by the relevant authorities. For example, according to the Ash Development Association of Australia (ADAA), NSW coal ash contains toxic elements that are in excess of regulatory guidelines. In particular, NSW coal ash typically comprises up to 31 different elements, at least 11 of which are classified as toxic according to the EPA Coal Ash Order of 2014. Gold and silver have also been found in trace amounts in some coal ash deposits.

[005] It is to be understood that, if any prior art is referred to herein, such reference does not constitute an admission that the prior art forms a part of the common general knowledge in the art, in Australia or any other country.

SUMMARY

[006] Disclosed herein is a process for the remediation of impounded coal ash. The process may be used to remediate coal ash that is impounded in a pond, such as may be located in the vicinity of a coal-fired power station. The coal ash may be impounded in an active pond or may be impounded in an inactive pond. The process may also be employed to treat coal ash prior to the coal ash being sent to a pond.

[007] The process as disclosed herein may remediate the impounded coal ash such that it may render the coal ash non-toxic. For example, each of the above-referenced toxic components may be recovered and/or encapsulated into a stable form, for either safe disposal or for sale to an end-user.

[008] For example, the process can comprise subjecting the coal ash to a magnetic separation treatment in which magnetisable material present in the coal ash is separated from non-magnetisable material present in the coal ash. In the case where the coal ash is held in a storage pond, the coal ash may first be removed from the storage pond. The coal ash removed from the storage pond may then be subjected to the magnetic separation treatment. In the case where the coal ash is not held in a storage pond (e.g. it is stored in a stockpile or heap), the coal ash may be subjected to the magnetic separation treatment directly. As set forth below, this step may advantageously produce a magnetic product (e.g. a magnetic component which may be sold e.g. as an exotic metal alloy mix).

[009] The process can also comprise passing the separated non-magnetisable coal ash to a water solubles separation treatment stage. In this stage, water-soluble components in the coal ash can be separated from non-water-soluble components in the coal ash.

Advantageously, the water-soluble components can be recovered and sold as valuable byproducts and/or used as components of cementitious products, as set forth below.

[010] The process can further comprise passing the resultant coal ash substantially free of non-magnetisable material and water-soluble components to an alumina recovery stage.

The alumina recovery stage can comprise one or more process stages by which a solid substantially comprising alumina is produced and is separated from a coal ash residue. Because the process can be configured and operated to reduce contaminants (e.g. unburnt carbon, organic matter, metals etc.) prior to the alumina recovery stage, the resultant recovered alumina may constitute a ‘compliant’ material. In this regard, the recovered alumina may be freely sold, e.g. for aluminium production or for use in ceramics, etc.

[011] The process as disclosed herein is able to be deployed to remediate impounded/stored coal ash which may otherwise be unusable. Through such remediation, the process can produce one or more useful and non-hazardous products. Also, because the impounded/stored coal ash is removed from storage, the coal ash pond will eventually be emptied of the coal ash (i.e. once all the impounded/stored coal as has been remediated). However, because water taken from the pond in the process can be recycled back to the pond, the remediated pond will continue to contain water and thus may be able to support aquatic life, etc.

[012] Advantageously, the process as disclosed herein aims to provide a comprehensive remediation process for coal ash which can account for most or all of the most common elements identified in NSW coal ash (based on analyses performed by the ADAA of coal ash from the Western Coalfields deposit and the Hunter Valley). The present process can complement and build on the process disclosed in AU 2021902692. In this regard, the process of AU 2021902692 provides a method of recovering impounded coal ash as one or more coal ash products (e.g. granular fill material, bulk fill material, supplementary cementitious material, aggregated ash products etc.). However, the process of AU 2021902692 does not provide a means of recovering the alumina in the coal ash, nor does it provide a means of stabilising the toxic elements present in the coal ash. Further, the process of AU 2021902692 does not provide a means of recovering, as valuable byproducts, the other metals, non-metals and metalloids present in the coal ash. The process disclosed herein may be used to remediate coal ash comprising the most common elements, providing a means for recovering these in a form for safe disposal or, for those of economic value, as saleable by-products.

[013] As outlined above, the present process enables recovery and treatment of the coal ash to produce alumina, which is the most valuable commodity found in coal ash on a commercial scale. It is also the second-most abundant component of coal ash, comprising approximately 24% by weight of NSW coal ash (the most abundant being silica at approximately 68% by weight). However, the process also allows toxic elements, which are present in trace amounts (approximately 100 ppm or less) and are embedded in the coal ash, to be recovered and converted to stable forms. For example, the toxic elements may be incorporated into cementitious products. Advantageously, the alumina product can be free from such contaminants.

[014] The process allows a range of other valuable products to be recovered at the same time (e.g. compounds commonly used in the manufacture of metal alloys). Such products may be ‘tailored’ to comply with customer specifications, as well as with Australian (or other relevant) Standards and relevant Environmental Protection Authority regulations.

[015] To minimise risks to operators and to ensure quality control, the process as disclosed herein may adopt modern process control, automation and monitoring techniques. This may include using an extensive Supervisory Control and Data Acquisition (SCADA) system coupled to in-line x-ray fluorescence (XRF) spectrometers to perform element analyses in close to real-time.

[016] In some embodiments, the one or more process stages of the alumina recovery stage may additionally produce a solid substantially comprising dolomite and calcite. The solid substantially comprising dolomite and calcite may be separated from the coal ash residue and alumina. The solid substantially comprising dolomite and calcite has several uses, depending on the relative ratios of magnesium and calcium. For example, if the solid is substantially dolomite, it can be used in the production of float glass. However, if the solid is substantially calcite, it can be used as a construction material.

Water Solubles Recovery

[017] In some embodiments, the water solubles separation stage may comprise separating an aqueous solution comprising the separated water-soluble components from a sludge residue comprising the non-water-soluble components. Depending on the characteristics of the impounded coal ash, the sludge residue comprising the non-water-soluble components may be a still-wet sludge-like fluid or a semi-solid or a residual slurry. It will be appreciated that the characteristics (e.g. filterability, viscosity etc.) of the sludge residue will determine how the aqueous solution comprising the water-soluble components may be separated from the sludge residue. For example, when the sludge residue has good filterability, said separation may be by filtration. However, when the sludge residue does not have good filterability, said separation may be by centrifugation. As a further example, the said separation may be by settlement, e.g. when the sludge residue tends to agglomerate. The type of separation technique employed can be determined based on the characteristics of the particular coal ash and its resultant sludge residue to be treated.

[018] In some embodiments, the aqueous solution comprising the separated water-soluble components may be subjected to one or more processing stages. In the one or more processing stages, one or more of the water-soluble elements may be recovered as a product.

[019] In some embodiments, the one or more processing stages may each comprise ionexchange. Each ion-exchange stage may be operated so as to recover one or more of the one or more water-soluble elements from the aqueous solution, for example, by selecting a particular ion-exchange resin. Advantageously, the eluant(s) used to regenerate the resin in each ion-exchange stage may be prepared from recycled saline process water recovered from other stages within the process. [020] In some embodiments, a final stage of the one or more ion-exchange stages may produce a solution comprising sodium. The solution comprising sodium may be recovered for use in a chlor-alkali plant.

[021] In some embodiments, the aqueous solution comprising the separated water-soluble components may pass through up to seven processing stages. Each processing stage may produce a respective metal or metalloid compound.

Alumina Recovery

[022] As above, the present process may be configured and operated to recover alumina as a ‘compliant’ material.

[023] For example, in some embodiments, the alumina recovery stage may comprise an acid solubles separation treatment. In the acid solubles separation treatment, acid-soluble components in the coal ash may be separated from a residue comprising non-acid-soluble components in the coal ash.

[024] In some embodiments, the acid solubles separation treatment may comprise subjecting the coal ash to acidic reaction conditions under which the acid-soluble components are leached therefrom.

[025] In some embodiments, the acidic reaction conditions may comprise adding hydrochloric acid to the coal ash. When the acidic reaction conditions comprise adding hydrochloric acid to the coal ash, one or more of the non-acid-soluble components may form a chloride precipitate. This is because many metal oxides react with hydrochloric acid to form metal chlorides, some of which are insoluble. The metal chlorides of the non-acid- soluble components that are insoluble in acid may then reprecipitate as a chloride precipitate.

[026] In some embodiments, the acid solubles separation treatment may further comprise separating a residue comprising the chloride precipitate and other non-acid-soluble components from a solution comprising the acid-soluble components. It will be appreciated that the characteristics (e.g. filterability, viscosity etc.) of the residue comprising the chloride precipitate and other non-acid-soluble components will determine how said sludge is separated.

Acid Solubles Recovery

[027] As part of producing a compliant (ready- to- use) alumina product, in some embodiments, the solution comprising the acid-soluble components (i.e. including Al) may be subjected to one or more further processing stages. In these one or more further processing stages, one or more of the acid-soluble components may be recovered as a product.

[028] In some embodiments, the one or more further processing stages each comprise ionexchange. In these embodiments, the use of hydrochloric acid in the acid solubles separation treatment is particularly advantageous because the presence of hydrochloric acid in the solution comprising the acid-soluble components does not adversely affect the performance of the resin in the ion-exchange compared to e.g. nitric acid. Oxidizing agents such as nitric acid are known to attack organic resins, resulting in resin degradation and potentially violent exothermic reactions (e.g. explosions), thereby posing a safety risk. Of further advantage, is that the eluant(s) used to regenerate the resin in each ion-exchange stage may be prepared from recycled saline process water recovered from other stages within the process, with the chloride forming part of such saline process water.

[029] In some embodiments, there may be multiple, for example up to five, processing (e.g. ion-exchange) stages. Each processing stage may produce a respective cation of a metal or a metalloid.

[030] In some embodiments, a resultant solution substantially free of the acid-soluble components may be recovered from a final stage of the one or more further processing stages and may be passed to a precipitation stage.

[031] In some embodiments, the precipitation stage may comprise adding sodium carbonate to the solution substantially free of the acid-soluble components. The sodium carbonate causes alumina to precipitate, and reacts to produce calcium carbonate and magnesium carbonate from the solution. Carbon dioxide produced due to the precipitation of alumina may be recovered and used to generate the sodium carbonate, thereby minimising the carbon footprint of the process. For example, the carbon dioxide can be passed (e.g. bubbled) through a solution comprising sodium hydroxide.

[032] In some embodiments, the precipitation stage may further comprise adding acid to a resultant slurry comprising the precipitated solid comprising alumina, calcium carbonate and magnesium carbonate. The acid may be added so as to redissolve calcium carbonate and magnesium carbonate therefrom. The addition of the acid may also be controlled such that the alumina remains as a solid precipitate. Carbon dioxide produced due to the dissolution of calcium carbonate and magnesium carbonate may be recovered and used to generate the sodium carbonate (as above), thereby minimising the carbon footprint of the process.

[033] In some embodiments, the acid may comprise hydrochloric acid. The use of hydrochloric acid is advantageous because the neutralisation reaction with sodium carbonate and the dissolution reaction with calcium carbonate/magnesium carbonate results in the formation of sodium chloride. Solution(s) recovered from the precipitation stage are thus saline (i.e. comprising sodium chloride) and may be suitable for regenerating certain resin(s) in the ion-exchange stages. Because the solution comprises sodium chloride, it may also suitable for use in a chlor-alkali plant.

[034] In some embodiments, the precipitated solid substantially comprising alumina may be separated from a resultant solution comprising dissolved Ca and Mg.

[035] In some embodiments, the solid substantially comprising alumina may be subjected to a thermal treatment. In the thermal treatment, excess water may be removed therefrom, thereby producing a dried substantially alumina product. Because of the processing stages as outlined above, the alumina product may be compliant and ‘ready-to-use’.

[036] In some embodiments, the solution comprising Ca and Mg may be recovered. This recovered solution may be passed to a further precipitation stage. In the further precipitation stage, sodium carbonate may be added thereto so as to reprecipitate calcium carbonate and magnesium carbonate in the form of dolomite and calcite.

[037] In some embodiments, the precipitated solid substantially comprising dolomite and calcite may be separated from a resultant solution substantially free of Ca and Mg. This latter solution may be recovered for reuse within the process. Because this solution may comprise sodium carbonate, it can be sent to a recovery plant where hydrochloric acid is added, thereby producing a sodium chloride solution and carbon dioxide. The carbon dioxide may be captured and used to generate more sodium carbonate for use in the precipitation of alumina, calcium carbonate and/or magnesium carbonate.

[038] In some embodiments, the solid substantially comprising dolomite and calcite may be subjected to thermal treatment. In a thermal treatment stage, excess water may be removed therefrom, thereby producing a dried substantially dolomite and calcite product. The resultant dried solid substantially comprising dolomite and calcite may be used as a component of a cementitious product or as a component of a structural light weight aggregate.

Coal Ash Pre-treatment

[039] In some embodiments, prior to the magnetic separation treatment, the coal ash may be mixed with water to form a coal ash slurry. The coal ash slurry may then be passed to the magnetic separation treatment. [040] In some embodiments, prior to the magnetic separation treatment, the coal ash may be passed to a flotation stage. In the flotation stage, unburnt carbon, asbestos fibres and organic matter (when present) may be separated as an overflow, with an underflow from the flotation stage being passed to the magnetic separation treatment. For example, unburnt carbon can be present in the coal ash as a result of incomplete combustion of coal in coal burners of the thermal power station. Asbestos fibres can be present in the coal ash as a result of construction materials used in the power station. Organic matter may also be present as a result of dredging the coal ash from e.g. an active storage pond. The unburnt carbon, asbestos fibres and organic matter may be recycled as a kind of biomass for reburning in suitable burners at the thermal power station. As the unburnt carbon, asbestos fibres and organic matter are combusted at the thermal power station, asbestos fibres will melt and be reduced to their elemental oxide form, thus eliminating the fibrous format which is a hazard to humans and the environment.

[041] In some embodiments, the separated magnetisable material may be recovered as a metal alloy mix. For example, the metal alloy mix may be recovered when the composition of the separated magnetisable material makes it valuable for use as an exotic metal alloy mix that may be sold to specialist steel foundries. However, in other embodiments, the separated magnetisable material may be recovered as individual elements. For example, when the composition of the separated magnetisable material is not suitable for use as an exotic metal alloy mix.

Non- Acid-Solubles Recovery

[042] In some embodiments, the residue comprising the chloride precipitate and other non- acid-soluble components separated from the solution comprising the acid-soluble components may be subjected to further treatment stages. In the further treatment stages, one or more of the non-acid-soluble components may be recovered as a respective product.

[043] In some embodiments, the further treatment stages may comprise mixing the residue comprising the chloride precipitate and other non-acid-soluble components with a hot weak alkaline solution followed by hot water. As the residue comprising the chloride precipitate and other non-acid-soluble components is mixed with a hot weak alkaline solution followed by hot water, the water-soluble chlorides may be caused to dissolve therefrom. The hot weak alkaline solution may be added so as to neutralise any acid present in the residue.

[044] In some embodiments, the further treatment stages may also comprise separating a solution comprising the water-soluble chlorides from a residual sludge comprising non-water- soluble chloride precipitates and other non-acid-soluble components. It will be appreciated that the characteristics (e.g. filterability, viscosity etc.) of the residual sludge comprising non- water-soluble chloride precipitates will determine how the solution comprising the water- soluble chlorides is separated therefrom.

[045] In some embodiments, the solution comprising the water-soluble chlorides may be subjected to one or more ion-exchange stages. In the one or more ion-exchange stages, one or more of the water-soluble chlorides may be recovered from the solution.

[046] In some embodiments, there may be up to two ion-exchange stages. Each ionexchange stage may produce a respective cation of a metal. The respective cation of a metal may be recovered from its respective ion-exchange stage as a concentrated eluate from backwashing the ion-exchange resin. Advantageously, the eluent used to backwash the ion-exchange resin may be a solution comprising sodium chloride recovered from elsewhere in the process.

[047] In some embodiments, a solution substantially free of water-soluble chlorides may be produced from a final stage of the one or more ion-exchange stages. Said solution may be recovered for use as process water.

[048] In some embodiments, one or more of the recovered water-soluble chlorides may be used as a component of a structural light weight aggregate or encapsulated in a cementitious product. For example, the cementitious product may be cement roof tiles, masonry blocks and/or concrete panels.

[049] In some embodiments, the hot water-soluble chlorides produced by these one or more ion-exchange stages may comprise Pb and/or Hg. Both Pb and Hg are toxic elements. Advantageously, by employing the present process to remediate impounded coal ash, the Pb and/or Hg may be converted for use as a component of a cementitious product. Once part of the cementitious product, the toxic elements may be trapped within the structure of the cementitious product and thus stabilised.

[050] In some embodiments, the residual sludge comprising non-water-soluble chloride precipitates and other non-acid-soluble components separated from the solution comprising the water-soluble chlorides may be recovered. Said recovered residual sludge may be mixed with a solution comprising sodium thiosulphate. As the residual sludge is mixed with a solution comprising sodium thiosulphate, precious metals (when present) may be dissolved therefrom.

[051] In some embodiments, a solution comprising the precious metals may be separated from a residual sludge substantially free of the precious metals. The precious metals in the solution may be extracted to produce one or more precious metal products. For example, the precious metals may be extracted by electrolysis.

[052] In some embodiments, the precious metals may comprise Au and/or Ag. In these embodiments, the Au and/or Ag is typically present in trace quantities.

[053] In some embodiments, a solution substantially free of the precious metals, such as a spent liquor, may be recovered for reuse within the process. For example, this solution may be recycled to produce the solution comprising sodium thiosulphate.

[054] In some embodiments, a residual sludge substantially free of the precious metals separated from the solution comprising the precious metals may be recovered. The residual sludge substantially free of the precious metals may be mixed with nitric acid. Said mixing of the residual sludge and nitric acid may cause components soluble in nitric acid to dissolve therefrom.

[055] In some embodiments, a solution comprising nitric acid-soluble components may be separated from a residual sludge. The solution comprising nitric acid-soluble components may be neutralised. For example, the solution comprising nitric acid-soluble components may be neutralised by the addition of an alkali solution thereto (e.g. sodium hydroxide). This may cause the nitric-acid soluble components to reprecipitate therefrom.

[056] In some embodiments, a resultant reprecipitated solid comprising the nitric-acid soluble components may be separated from a resultant neutralised solution. For example, the reprecipitated solid may be separated from the neutralised solution by centrifugation. The neutralised solution may optionally be recovered for reuse in the process. For example, the neutralised solution may be recycled back to the nitric acid preparation area and may be used to produce the nitric acid.

[057] In some embodiments, the separated reprecipitated solid comprising the nitric-acid soluble components may be subjected to thermal treatment. In the thermal treatment, the separated reprecipitated solid may be dried. Also, the nitric-acid soluble components may be substantially converted to oxides.

[058] In some embodiments, the components soluble in nitric acid may comprise Ti. Advantageously, the Ti may be converted to TiC>2 during the thermal treatment, which is used in a variety of commercial and industrial products.

[059] In some embodiments, the residual sludge separated from the solution comprising nitric acid-soluble components may be recovered. The residual sludge may be mixed with a dilute alkali solution. By mixing the residual sludge with a dilute alkali solution, dilute-alkali- soluble components may be selectively dissolved therefrom.

[060] In some embodiments, a resultant solution comprising the dilute-alkali-soluble components may be separated from a residual sludge comprising non-dilute-alkali-soluble components.

[061] In some embodiments, one or more of the dilute-alkali-soluble components in the solution may be extracted. For example, one or more of the dilute-alkali-soluble components may be extracted by ion-exchange.

[062] In some embodiments, the extracted dilute-alkali-soluble components may be added to the separated magnetisable material. For example, such addition may be made to enhance the properties of the separated magnetisable material.

[063] In some embodiments, the remaining solution substantially free of the dilute-alkali- soluble components may be recovered for reuse in the process. For example, this solution may be used to produce the dilute alkali solution.

[064] In some embodiments, the dilute-alkali-soluble components may comprise Mo. In these embodiments, it is particularly advantageous to add the extracted Mo to the separated magnetisable material so that a highly valuable exotic metal alloy can be produced. For example, the exotic metal alloys Hastelloy, Inconel, Martensitic Stainless Steel and Austenitic Stainless Steel all comprise Mo in various quantities.

[065] In some embodiments, the dilute alkali solution mixed with the residual sludge may comprise sodium hydroxide at a concentration of about 1M. Advantageously, the sodium hydroxide may be produced from a chlor-alkali plant (as referred to above and as described hereafter).

[066] In some embodiments, the residual sludge comprising non-dilute-alkali-soluble components that is separated from the solution comprising the dilute-alkali-soluble components may be recovered. This recovered residual sludge may be mixed with a concentrated alkali solution so as to selectively dissolve concentrated-alkali-soluble components therefrom. The concentrated alkali solution mixed with the residual sludge may comprise sodium hydroxide at a concentration of about 18M. Advantageously, the sodium hydroxide may be produced from the chlor-alkali plant.

[067] In some embodiments, a resultant concentrated alkali solution comprising the concentrated-alkali-soluble components may be separated from a residual sludge. [068] In some embodiments, acid may be added to the concentrated alkali solution. Because acid is added, the concentrated alkali solution may be neutralised, and the concentrated-alkali-soluble components may be caused to precipitate therefrom. For example, the acid may be hydrochloric acid.

[069] In some embodiments, the precipitate comprising the concentrated-alkali-soluble components may be separated from the neutralised solution. The neutralised solution may optionally be recovered for reuse within the process. For example, in the embodiments where the acid is hydrochloric acid, the neutralised solution comprises sodium chloride and may be used in the chlor-alkali plant or in the preparation of ion-exchange regeneration solutions.

[070] In some embodiments, the concentrated-alkali-soluble component may comprise Ge. Typically, recovered Ge will be in the form of GeO2, which can be used in the production of certain semiconductor materials and elemental Ge.

[071] In some embodiments, the residual sludge separated from the concentrated alkali solution comprising the concentrated-alkali-soluble components may be recovered. The residual sludge may be subjected to dewatering such that excess water is removed therefrom. A residual solid substantially free of liquid may thereby be produced.

[072] In some embodiments, the residual solid substantially free of liquid may substantially comprise Si and Zr. The residual solid substantially free of liquid may therefore be an essentially inert solid.

[073] In some embodiments, the residual solid substantially free of liquid may be used as an earth fill material. However, in other embodiments, the solid substantially free of liquid may be subjected to thermal treatment in which remaining water is removed therefrom, thereby producing a dried solid product. In these other embodiments, the dried solid product may be used for example, as a component for making a structural light weight aggregate.

Components of the Coal Ash

[074] In some embodiments, the magnetisable material in the coal ash may comprise one or more of: Cr, Cu, Co, Fe, Mn, Ni, W, V.

[075] In some embodiments, the water-soluble components in the coal ash may comprise one or more of: As, Ba, B, Li, K, Se, Na.

[076] In some embodiments, the acid-soluble components in the coal ash may comprise one or more of: Sb, Be, Cd, Sn, Zn. [077] Thus, in some embodiments, the coal ash may comprise for example 33 elements: Ge, Pb, Hg, Mo, Si, Ti, Zr, Cr, Cu, Co, Fe, Mn, Ni, W, V, As, Ba, B, Li, K, Se, Na, Sb, Be, Cd, Sn, Zn, Al, Ca, Mg, as well as Au, Ag (in trace quantities) and carbon.

[078] In some embodiments, the solid substantially comprising dolomite and calcite may be separated into a high magnesium product and a low magnesium product. However, in other embodiments, the solid substantially comprising dolomite and calcite may, as outlined above, be used as a component of a cementitious product. In these other embodiments, the cementitious product may comprise a structural light weight aggregate.

Chlor-Alkali Production

[079] In some embodiments, the process may comprise a chlor-alkali plant. In the chloralkali plant, a near-saturated solution of sodium chloride may be used to produce hydrogen gas, chloride gas and sodium hydroxide. Sodium hydroxide may be used as a reagent within the process. For example, sodium hydroxide may be used to recover the non-acid-soluble components of the coal ash. Hydrochloric acid may be produced from the hydrogen gas and chloride gas and may be used to recover the acid-soluble components of the coal ash. The chlor-alkali plant may also produce a solution comprising sodium chloride with a lower concentration of sodium chloride compared to the near-saturated solution of sodium chloride. Hydrochloric acid, sodium chloride or sodium hydroxide may be suitable to regenerate resin in the ion-exchange stages, depending on the resin selected.

[080] In some embodiments, substantially renewable energy may be used to generate electricity for powering the process. This may reduce the carbon footprint and increase the sustainability of the process.

[081] In some embodiments, carbon dioxide produced by the process may be recovered. The recovered carbon dioxide may be used in the generation of a carbonate solution for use in the production of the solid substantially comprising dolomite and calcite that is separated from the coal ash residue and alumina in the alumina recovery stage.

[082] In some embodiments, substantially all steam generated by the process may be condensed as process water for reuse in the process. This may reduce the demand for fresh water, thereby reducing the environmental impact of the process.

[083] Further aspects of the invention are as set out in the drawings and description, and in the claims. BRIEF DESCRIPTION OF THE DRAWINGS

[084] Embodiments will now be described by way of example only, with reference to the accompanying drawings in which:

[085] Fig. 1 is a concept flow diagram, set out in simple block diagram form, of an embodiment of a process for remediating coal ash as disclosed herein.

[086] Fig. 2 is a block diagram of the initial process stages of the coal ash remediation process of Fig. 1.

[087] Fig. 2a is a schematic illustrating the operation of a rotary drum magnetic separator suitable for use in the process, Fig. 2b is a schematic illustrating the operation of a dewatering press suitable for use in the process, Fig. 2c is a diagram of an indirectly-heated rotary dryer for use in the process, and Fig. 2d shows how the vanes inside such a rotary dryer can be constructed.

[088] Fig. 3 is a block diagram of water-soluble components recovery stages of the process of Fig. 1.

[089] Fig. 3a is a schematic illustrating the operation of a single ion-exchange column suitable for use in a batch ion-exchange stage and Fig. 3b is a schematic illustrating the operation of a continuous or semi-continuous ion-exchange process.

[090] Fig. 4 is a block diagram of the acid leaching stage of the coal ash remediation process of Fig. 1.

[091] Fig. 5 is a block diagram of a first part of non-acid-soluble components recovery stages of the coal ash remediation process of Fig. 1.

[092] Fig. 6 is a block diagram of a second part the non-acid-soluble components recovery stages of the coal ash remediation process of Fig. 1.

[093] Fig. 7 is a block diagram of an acid-soluble components recovery stage of the coal ash remediation process of Fig. 1.

[094] Fig. 8 is a block diagram of alumina, and dolomite/calcite precipitation stages of the coal ash remediation process of Fig. 1.

[095] Fig. 9 is a block diagram of a process by which nitric acid is regenerated for use in the non-acid-soluble components recovery area of Fig. 5.

[096] Fig. 10 is a schematic of a chlor-alkali plant, such as may form part of the auxiliary equipment of the coal ash remediation process of Fig. 1. DETAILED DESCRIPTION OF SPECIFIC EMBODIMENTS

[097] In the following detailed description, reference is made to accompanying drawings which form a part of the detailed description. The illustrative embodiments described in the detailed description, depicted in the drawings and defined in the claims, are not intended to be limiting. Other embodiments may be utilised and other changes may be made without departing from the spirit or scope of the subject matter presented. It will be readily understood that the aspects of the present disclosure, as generally described herein and illustrated in the drawings can be arranged, substituted, combined, separated and designed in a wide variety of different configurations, all of which are contemplated in this disclosure.

[098] The following description discloses an embodiment of a process for remediating coal ash. The coal ash can be produced by burning coal in burners for boilers of e.g. a power station, or in other industrial processes that burn coal. By remediating the coal ash, an alumina product can be produced. Other by-products can also be produced in varying proportions and with varying compositions. The proportions and compositions of certain byproducts can be selected based on the composition of the coal ash to be remediated along with economic viability of producing the various by-products. The remediation process is also designed and operated towards full remediation of coal ash comprising the 30 or so most common elements found in coal ash in NSW. The remediation process is also designed and operated such that the toxic components of the coal ash can be recovered, and such that the products can comply with Australian Standards and EPA regulations to thereby be certified by environmental authorities.

[099] The process for remediating coal ash as disclosed herein may be used in conjunction with the process of the inventor’s co-pending application AU 2021902692 with the latter process producing up to four cementitious products one or more of which may be free of toxic elements. The process as disclosed herein can produce additional saleable products. Such products comprise components of the impounded coal ash and are removed as part of this process, with a number being recovered as valuable products, e.g. alumina, dolomite/calcite, gold, silver, and various other metals and metalloids, etc.

[100] Referring now to Fig. 1 , a concept flow diagram, set out in block diagram form, illustrates a process 10 for remediating coal ash to recover, inter alia, alumina. The total process 10 is divided into what may be considered eight ‘stages’, as follows: a. Impounded pond ash removal b. Material grading c. Magnetic separation d. Water-soluble component recovery e. Acid leaching f. Non-acid-soluble component recovery g. Acid-soluble component recovery h. Precipitation

[101] The feed to process 10 is impounded coal ash from a storage pond (omitted in Fig. 1 but shown in Fig. 2). According to the Ash Development Association of Australia (ADAA), impounded coal ash from NSW coal-fired power stations primarily comprises the following elements: Al, Sb, As, Ba, Be, B, Cd, Ca, Cr, Co, Cu, Ge, Fe, Pb, Li, Mg, Mn, Hg, Mo, Ni, K, Se, Si, Na, Ti, Sn, W, V, Zn, Zr. Additionally, the coal ash will contain a level of unburnt carbon, C (as measured by LOI) and may contain trace amounts of Au and Ag. Thus, the coal ash feed to the process 10 comprises a mixture of many different elements. It can be seen as advantageous to develop a remediation process in which each of the most common elements is accounted for and, furthermore, wherein each of the most common elements can be recovered in the form of either a stabilised product for safe disposal or as a valuable by-product for sale.

[102] The elements of the coal ash considered to be most toxic (due to their inherent toxicity and/or concentration in the coal ash) are: As, B, Cd, Cr, Cu, Pb, Hg, Mo, Ni, Se, Zn. It is of further advantage to develop a remediation process in which each of these toxic elements is recovered and captured in a stable form.

[103] After combustion of coal in the presence of oxygen (i.e. from air), such as in a power station, the elements comprising the coal ash typically reside in the form of oxides. Once the coal ash is placed under water (e.g. as impounded ash in a storage pond), some conversion of oxides to hydroxides can occur. The properties of the impounded coal ash fed to the process 10 are described in greater detail below with reference to Example 1.

[104] The process 10 allows aluminium to be recovered in a saleable form (i.e. as alumina). This is particularly advantageous because alumina is currently the component of coal ash that is of greatest value, due to the high aluminium content and its market price. Of further advantage is that the toxic elements are stabilised for safe disposal or for incorporation into valuable by-products (e.g. cementitious products, such as those produced in co-pending AU 2021902692). Other elements (i.e. the non-aluminium and non-toxic elements) present can likewise be recovered as valuable by-products. By recovering the majority of the components of the coal ash as valuable by-products, the process becomes economically viable - i.e. because sale of these by-products helps to offset the capital/operating costs of a coal ash processing plant.

[105] To the best of the inventor’s knowledge, the process 10 is the first known process which aims to fully remediate coal ash, with all the main elements of coal ash being accounted for in the process 10.

[106] The initial stages of the process 10, including a flotation stage 200, and a magnetic separation treatment 300, are described in more detail below with reference to Fig. 2.

However, in general, to produce the feed for the process 10, coal ash slurry is removed from a pond 102 (Fig. 2), for example by a hydraulic dredge or by excavator. If the coal ash is not held in a storage pond but is e.g. held in a stockpile or an inactive pond, a slurry 108 can be produced from the coal ash stockpile or the inactive pond by adding water to enable it to be treated by the process 10.

[107] The slurry dredged from the pond by the dredge (or that is produced from stockpiled coal ash or inactive pond) is pumped to either to an optional trash screen 110 (Fig. 2), or directly to a material grading area 200. The optional trash screen 110 may be used when the coal ash comprises a significant portion of vegetative waste, such as when the coal ash originates from an active pond.

[108] When the slurry is pumped to the material grading area 200, the slurry is pumped to a flotation stage 202. In the flotation stage 202, flotation is performed in a flotation cell to separate carbon, organic matter (when present) and other contaminants, such as fibrous asbestos, from the fines slurry. Organic matter (e.g. plant material, pond life, etc.) can be present as a result of dredging the coal ash from a storage pond. The organic matter floats to the top of the flotation cell and is collected as an overflow 204. As explained below, the overflow comprising the separated carbon, organic matter and other contaminants 204 can be recycled to a power station, etc. as a kind of biomass for combustion in the boiler burners. As also explained below, if the coal ash does not comprise a significant amount of carbon, organic matter and/or other contaminants, the flotation stage 202 may be bypassed.

[109] The fines slurry collects at the bottom of the flotation cell 202, forming an underflow 206. The underflow 206 is pumped to a coarse filter where fines are separated from coarser material present in the slurry. The coarser material 210 is collected. The coarser material 210 typically comprises sintered bottom ash from the boilers and is used to form a granular fill material product. Fines present in the slurry pass through the coarse filter to form a pre- treated ash slurry 212, which is pumped by slurry pump to the magnetic separation stage 300.

[110] In the magnetic separation treatment 300, magnetisable material is separated from non-magnetisable material 304 present in the treated ash slurry 212. The magnetisable material is collected as a sludge 306 and is dewatered 308 and dried 314 to form an exotic metal alloy product mix 320 (or it can be further processed to produce individual metal products).

[111] The non-magnetisable material 304 collected in the magnetic separation treatment 300 is passed to a water solubles recovery area 400 (Fig. 3). In the water solubles recovery area 400, an aqueous solution 406 comprising the water-soluble components 402 is separated from a solid 404 comprising the non-water-soluble components in the non- magnetisable material (i.e. an ash sludge), such as by filtration. As described below, the ash sludge 404 is passed to acid leaching stage 500, whereas the aqueous solution 406 comprising the water-soluble components 402 is subjected to one or more processing stages in which one or more products comprising the respective water-soluble components is produced. The one or more processing stages typically each comprise ion-exchange, with resin(s) selected based on the water-soluble component(s) to be recovered. The water solubles recovery area 400 is described in more detail below with reference to Fig. 3.

[112] The solid 404 comprising the non-water-soluble components from the water solubles recovery area 400 is recovered as a slurry/sludge 404. The recovered slurry is pumped by a slurry pump to the acid leaching stage 500. In the acid leaching stage 500, acid is added to the slurry and the mixture is subjected to reaction conditions under which acid-soluble components are dissolved from the solids in the slurry. Typically, the acid is hydrochloric acid 504 because this concomitantly promotes the formation of metal chlorides, a small number of which are insoluble in the acid and will form a precipitate. A solution 502 comprising the acid-soluble components is separated from a solid 505 comprising the non- acid-soluble components (which solid also includes the precipitated chlorides), such as by filtration. These processing stages are described in more detail below with reference to Fig. 4.

[113] The solution 502 comprising the acid-soluble components is passed to an acidsolubles recovery area 700. In the acid-solubles recovery area 700, the solution comprising the acid-soluble components 502 is subjected to a number of different processing stages by which one or more products 702 comprising the acid-soluble components are recovered. The processing stages typically comprise ion-exchange, with resin(s) selected for each stage based on the properties of the acid-insoluble component(s) to be recovered. The remaining solution 728 exiting the one or more processing stages of the acid solubles recovery area 700 comprises substantially Al, Ca and Mg. The acid-solubles recovery area 700 is described in more detail below with reference to Fig. 7.

[114] The solution 728 comprising substantially Al, Ca and Mg is passed from the acidsolubles recovery area 700 to a precipitation area 800. In the precipitation area 800, sodium carbonate 842 is added to the solution 728 so as to first precipitate a solid product comprising substantially alumina, calcite and dolomite. The calcite and dolomite are then redissolved by adding hydrochloric acid 812 to the slurry comprising the precipitate. The remaining solid comprises substantially alumina and is separated from the solution (which comprises Ca and Mg), thereby producing the alumina product 802.

[115] Sodium carbonate 842 is added to the remaining solution so as to reprecipitate Ca and Mg as a solid 804 substantially comprising dolomite and calcite. The precipitation area 800 is described in more detail below with reference to Fig. 8.

[116] The residue 505 comprising the non-acid-soluble components from the acid leaching area 500 is passed to a non-acid-solubles recovery area 600. In the non-acid-solubles recovery area 600, the residue is subjected to one or more processing stages by which one or more products comprising non-acid-soluble components 699 are produced. The one or more processing stages depend on the properties of the non-acid-soluble components to be recovered. Typically, the one or more processing stages comprise a combination of physical and/or chemical treatments. The non-acid-solubles recovery area 600 is described in more detail below with reference to Fig. 5.

[117] The process 10 further comprises a chlor-alkali plant 950 (Fig. 10). In the chlor-alkali plant 950, solution(s) comprising sodium chloride recovered from elsewhere in the process 10 are used to generate sodium hydroxide 968, hydrogen gas 966 and chlorine gas 964. The sodium hydroxide 968 is recovered for use in the process 10, for example, in the non- acid-soluble recovery area. The hydrogen gas 966 and chlorine gas 964 are likewise recovered and used to produce hydrochloric acid. Thus, the process 10 regenerates a number of its reagents.

[118] Where possible, renewable energy is used to generate electricity for powering the process 10. For example, a solar power plant can be constructed in the vicinity of the coal ash remediation plant. This increases the sustainability of the process 10. The sustainability of the process 10 is further increased by recovering process condensates (e.g. steam produced from the various dryers) for reuse as process water, recovering saline process water for reuse, and capturing emitted carbon dioxide, e.g. from the precipitation area 800, and reacting it to form further reagents.

[119] Specific details of the process 10 for accomplishing each of the operational stages, as outlined above, will now be described.

Impounded Pond Ash Removal - Fig. 2

[120] Referring now to Fig. 2, a flow diagram of the initial stages of process 10 is shown. The coal ash 104 is first removed from the storage pond 102 (e.g. by dredging and/or pumping). If required, water 106 is optionally added to the coal ash 104, i.e. to create a pumpable slurry 108.

[121] Coal ash waste, such as generated by coal combustion in a coal-fired power station, is often stored in ponds in proximity to the power station. If the pond is active, the pond will contain water to suppress the ash dust. Because the coal ash is typically denser than the pond liquid, the majority of ash will settle to the bottom of the pond over time. Once accumulated, it can be covered by organic matter growth, pond silt, etc.

[122] As in AU 2021902692, the settled coal ash is disturbed by mechanical means to generate a pumpable suspension. Typically, this involves the use of a hydraulic dredge installed on a floating hydraulic dredge barge. The dredge barge is deployed into the pond comprising the coal ash waste. The settled coal ash is mechanically disturbed by the dredge and is sucked in as a slurry by a pump installed on the dredge.

[123] When the pond 102 is an active pond, the removed coal ash 104 will be in the form of a slurry. As above, additional water 106 can be mixed with the pump on the dredge to increase the pumpability of the slurry 108. Should the slurry pump require a source of cleaner water than the disturbed pond water (e.g. for proper pump operation and/or to prevent fouling and blockage and/or minimise consumption of pond water), water extracted from a centrifuge bank 308 can be recycled back to the slurry pump directly.

[124] If the pond 102 is inactive, the pond will have been drained and capped with soil. To access the ash, the soil is removed to expose the ash. The ash is then removed from the pond, for example by using an excavator. In this case, the excavated ash 104 is mixed with water 106 to form a pumpable slurry 108 that can be easily transported. The water 106 can be a combination of water from the centrifuge bank 308 that is recycled back and/or other recycled process water and/or fresh water. [125] As the coal ash is removed from the pond and mixed with water to create a pumpable slurry 108, the components of the coal ash that are soluble in water tend to dissolve from the coal ash. Because the water-soluble components dissolve from the coal ash, the liquid component of the slurry 108 is a solution comprising the water-soluble components. The solid component of the slurry 108 is the remaining solid coal ash, now substantially free of the water-soluble components.

[126] The slurry 108 is pumped to a trash screen 110 in which non-ash waste, e.g. vegetative waste and solids, are removed from the slurry 108. This both maximises the quality of downstream products and protects/optimises downstream process equipment. The remaining slurry comprising ash material is collected as a clean ash slurry 114. The clean ash slurry 114 is pumped, e.g. by a slurry pump, to the flotation stage 202 in the separation area 200.

[127] The trash screen 110 can be in the form of a conveyor with perforations across which the slurry passes or in the form of a vertical grate. As the slurry 108 passes over/through the trash screen 110, the finer particles and liquids pass through the perforations, because they are smaller than the size of the perforations. However, the size of the perforations on the trash screen 110 (or the size of the slits on the grate) is preselected to generally prevent non-ash waste material and large solids, such as vegetative waste and tree roots, from passing therethrough. The non-ash waste materials 112 stopped from passing through the trash screen 110 collect on (or at) the trash screen 110. The non-ash waste material 112 typically comprises solids that are larger than the size of the perforations or grates of the trash screen 110. The non-ash waste material 112 can also comprise asbestos in concrete form and/or waste building materials that have been dumped into the pond 102. The non-ash waste 112 is collected and sent to a composting facility for disposal.

[128] The trash screen 110 is typically used when remediating impounded coal ash from an active pond. This is because, as the impounded coal ash is dredged from the active pond, non-ash material such as vegetative waste, tree roots, etc., becomes entrained with the coal ash. The coarse filter 110 may be by-passed when remediating impounded coal ash from an inactive pond because, once an overlying earth layer has been removed, the excavated coal ash tends to comprise less non-ash waste materials.

Material Grading Area 200 - Fig. 2 [129] In the material grading area 200, carbonaceous material and sintered bottom ash in the clean ash slurry 114 are removed, thereby producing a treated ash slurry 212 which is suitable to feed to the magnetic separation treatment stage 300.

[130] The clean ash slurry 114 is pumped by the slurry pump to the flotation stage 202. Pond coal ash usually contains unburnt carbon. Pond coal ash can also contain organic matter, for example, pond grass growing at and beneath the pond surface in the case of active ponds. In the case of inactive ponds, grass may be growing on the soil used to cap the pond. The pond coal ash can further comprise fibrous asbestos. Fibrous asbestos can originate from insulating material used on pipes etc., especially where pipes in the power station were lagged for insulation purposes with asbestos mixed with magnesia. In some cases, fibrous asbestos insulating materials were disposed of in the ash ponds. The flotation stage 202 is designed and operated to separate these further contaminants from a slurry comprising the lighter fines.

[131] In the flotation stage 202, flotation is carried out using one or more flotation cells 214 that are designed and operated using a froth flotation procedure. The type and function of the flotation cell is described in detail in AU 2021902692.

[132] The clean ash slurry 114 enters towards the top of the flotation cell. The flotation cell 214 comprises an agitator which acts to keep the fines in the slurry in suspension and also acts as a sparger through which pressurised air is caused to enter the flotation cell 214 as bubbles. A frothing agent is typically added to the cell 214 to promote frothing. The type and amount of frothing agent depends on the characteristics of the clean ash slurry 114 and is adjusted to ensure adequate recovery in the froth of unburnt carbon, organic matter (and asbestos, if present).

[133] In the cell, hydrophobic particles, such as unburnt carbon, organic matter and asbestos, attach to the air bubbles. Because the air bubbles (and attached hydrophobic particles) are less dense than the surrounding slurry, they rise to the surface, forming froth at the top of the flotation cell. The froth (comprising air and the hydrophobic particles) forms an overflow 204 from the flotation cell and is collected by means of a collection launder. In this way, the unburnt carbon, organic matter and asbestos are separated from the ash slurry.

[134] The overflow 204 comprising the unburnt carbon, organic matter and asbestos is eventually passed to the boilers of a nearby power station for combustion in the boiler burners as a biomass. During combustion, unburnt carbon and organic matter are incinerated and the fibrous asbestos is vaporised back to its basic elements. The vaporised asbestos is caught in bag filters attached to the power station gas outlet. [135] The carbon-free ash slurry 206 (now substantially free of unburnt carbon, organic matter and fibrous asbestos) is discharged as an underflow from the bottom of the flotation cell 214. The composition of the carbon-free ash slurry 206 is continuously monitored to ensure the fraction of unwanted materials reporting to the carbon-free ash slurry 206 is minimised. The type and amount of frothing agent(s) is adjusted accordingly, as is the residence time in the flotation stage, to increase the purity of the carbon-free ash slurry 206.

[136] The carbon-free ash slurry 206 is pumped by a slurry pump to a coarse filtration stage 208. In the coarse filtration stage 208, coarser material is removed (i.e. filtered) from the carbon-free ash slurry 206. Removal of coarser material both maximises the quality of downstream products and protects/optimises downstream process equipment.

[137] Typically, the filter will be an inclined static screen filter, i.e. of the type and function described in AU 2021902692. As the slurry passes to the static screen filter, the slurry flows down over the inclined screen surface of the screen filter which contains perforations. The size of the perforations on the screen surface is preselected to stop coarser material from passing therethrough. The coarser material 210 collects at the bottom of the inclined static screen filter. The coarser material 210 typically comprises sintered bottom ash (i.e. with diameters greater than the size of the perforations on the screen surface). The coarser material 210 is collected and can form a granular fill product (i.e. as described in AU 2021902692).

[138] Finer particles (i.e. with diameters smaller than the size of the perforations on the screen surface) and liquids pass through the perforations and collect as a treated ash slurry 212. The treated ash slurry 212 is pumped by a slurry pump to the magnetic separation stage 300.

Magnetic Separation Stage 300 - Fig. 2

[139] In the magnetic separation stage 300, the magnetisable material in the treated ash slurry 212 is separated from the non-magnetisable material in the treated ash slurry 212. The magnetisable material is separated out because it represents a valuable product of the process 10. The magnetisable material is separated using one or more drum magnetic separators 302 that are designed and operated to separate the magnetisable material 306 from the non-magnetisable material 304 in the treated ash slurry 206.

[140] Fig. 2a is a schematic that illustrates the internals of a drum magnetic separator 302 operating in a co-current motion. The treated ash slurry 206 is fed into a feed channel 302-1 of the drum magnetic separator 302. The drum magnetic separator 302 comprises a rotating shell 302-2. In the co-current arrangement illustrated in Fig. 2a, the rotating shell 302-2 rotates in the same direction R as the incoming feed 302-1. It will be appreciated that a counter-current rotation may also be employed in which the rotating shell 302-2 rotates in an opposite direction to the incoming feed 302-1.

[141] The rotating shell 302-2 has a permanent magnet 302-3 affixed therein, with a corresponding magnetic field. The permanent magnet 302-3 is stationary (fixed) with regard to the rotating shell 302-2. That is, the rotating shell 302-2 rotates outside the permanent magnet 302-3. Thus, the magnetic field due to the permanent magnet 302-3 likewise remains stationary. The magnet 302-3 extends from the feed channel 302-1 to a magnetisable material outlet channel 302-4.

[142] As the treated ash slurry 206 is fed into the feed channel 302-1 , the magnetisable components of the treated ash slurry 206 are attracted to the rotating shell 302-2, because of the magnetic field present due to the permanent magnet 302-3. The magnetisable components attach to the rotating shell 302-2 and rotate with the rotating shell 302-2, past a first outlet of the drum 302-5.

[143] Conversely, the non-magnetisable material is not attracted to the rotating shell 302-2 and thus flows (i.e. drops) off the drum at the first outlet 302-5. Typically, substantially all the liquid component of the treated ash slurry 206 also flows off the drum at the first outlet 302- 5. Thus, at outlet 302-5, a slurry 304 comprising the non-magnetisable material is collected.

[144] Because the magnetisable components are attached to the rotating shell 302-2, they rotate past the first outlet 302-5. The magnet 302-3 ends at a second (magnetisable material) outlet 302-4 of the drum 302. At the location where the magnet 302-3 ends, the magnetic components are no longer attracted to the rotating shell 302-2. The magnetic components detach from the rotating shell 302-2 and are collected at the outlet 302-4, forming the magnetic component of the ash slurry. The magnetic component is typically in the form of a sludge 306 of magnetisable material. The sludge 306 of magnetisable material is passed to a dewatering centrifuge 308.

[145] Referring to Fig. 2b, a schematic of the internals of a dewatering centrifuge 308 is shown. The sludge 306 is passed through the inlet 308-1 of the dewatering centrifuge 308 into the rotating bowl 308-2. The sludge 306 is accelerated into a feed section 308-6 of the bowl 308-2. The bowl 308-2 has a conical shape and rotates in the direction indicated by the arrow 308-3. Inside the bowl 308-2 is a rotating screw conveyor 308-5, which rotates in a direction opposite to direction 308-3. The bowl 308-2 is enclosed in an outer casing (not shown in Fig. 2b).

[146] As the sludge 306 is accelerated into the feed section 308-6 of the bowl 308-2, the rotating screw conveyor 308-5 acts to move the sludge through the centrifuge in direction 308-4, i.e. whilst the bowl 308-2 rotates. The rotation of the bowl 308-2 (and the screw conveyor 308-5) causes a centrifugal force to be exerted on the sludge. The centrifugal force causes the solids in the sludge to move to the outside edge 308-10 of the bowl 308-2, because the solids have more inertia compared to the liquids. Since the liquids 308-7 have comparatively less inertia, they tend to stay closer to the centre of rotation.

[147] As the screw conveyor 308-5 rotates, the plates 308-11 of the screw conveyor push the solids along the outside edge 308-10, moving them in the direction 308-4. The plates 308-11 of the screw conveyor 308-5 have small perforations through which the liquid 308-7 passes. This causes the liquid 308-7 to move counter-currently to the solids and collect at a centrate outlet 308-8. Meanwhile, due to the movement of the screw conveyor 308-5, the solids continue moving in direction 308-4, and collect at the beach 308-9. At the beach 308- 9, the solids are partially dried, before exiting at a solids outlet 308-10. It will be appreciated that the moisture content of the solids at the outlet 308-10 is dependent on many factors - e.g. moisture content of the sludge at the inlet 308-1, physical characteristics of the sludge, size of centrifuge, time in centrifuge, etc. Typically, the solids content at the outlet 308-10 is between about 40-60% dry solids.

[148] The solid 310 (now with a reduced moisture content) collected from the solids outlet of the centrifuge 308-10 is composed of magnetisable materials that are insoluble in water (i.e. an alloy of metallic compounds). In particular, the metal alloy solid 310 can comprise: FesC , C^Ch, COO, CUO, MnO2, NiO, WO2 V2O5.

[149] The liquor 312 collected from the centrate outlet of the centrifuge 308 is recycled back to the pond 102, when the pond 102 is active. When the pond 102 is inactive, the liquor 312 can comprise part of the water 106 used to form the coal ash slurry 108.

[150] The metal alloy solid 310 is transported, e.g. by means of a solids conveyor, to a drying stage 314. In the drying stage 314, the moisture content of the metal alloy solid 310 is further reduced, resulting in a dried metal alloy product. Typically, an indirectly heated rotary drum dryer 316 (see Figs. 2c and 2d) is used in the drying stage 314.

[151] In this regard, Fig. 2c shows a schematic of an indirectly-heated rotary drum dryer 316 and Fig. 2d shows the vanes inside such a dryer. The indirectly-heated rotary drum dryer 316 comprises a rotating drum 316-1 enclosed in a furnace 316-2. The furnace 316-2 of the rotary dryer is externally heated via burners 316-3.

[152] The metal alloy solid 310 is fed into a material inlet 316-4 of the rotary drum dryer 316. Typically, the material inlet 316-4 comprises a hopper and a screw conveyor which allows the feed material 310 to be fed at a controlled rate into the rotating drum 316-1 via the material inlet 316-4.

[153] Piping 316-5 for gas and air feeds the burners 316-3. Combustion of fuel (e.g. natural gas) in the burner 316-3 produces hot vapour which provides thermal energy/heat to the rotating drum 316-1 (i.e. the thermal energy produced from the combustion of the gas heats the rotating drum 316-1). The hot vapour comprises primarily carbon dioxide and water (i.e. from the combustion of fuel 316-5 in the burners 316-3), as well as smaller concentrations of carbon monoxide and carbon (i.e. from incomplete combustion). The hot vapour is released to the atmosphere as exhaust gases, via exhaust 316-6. Ideally, solar energy plants are installed in the vicinity of the coal ash ponds, allowing solar energy to be used as the thermal energy source for dryer 316. Using solar energy reduces the carbon footprint of the process 10, compared to using gas combustion.

[154] Fig. 2d shows vanes 316-1 A present inside the rotating drum 316-1. The vanes 316- 1A are attached to the internal walls of the rotating drum 316-1 and facilitate the continuous turning over and breaking down of the metal alloy solid into individual particles, as the metal alloy solid is thermally treated. The internal walls of the rotating drum 316-1 and the vanes 316-1A are heated by the furnace 316-2. The metal alloy solid makes contact with the heated internal walls and vanes as it progresses through the rotating drum 316-1, whereby the metal alloy solid is dried.

[155] Vapour is produced in the rotating drum 316-1 as the metal alloy solid is dried. Dried metal alloy solid product and vapours exit the rotating drum at an exit 316-7 (discharge breech). At the exit 316-7, a vapour stream is separated from the metal alloy product. The vapour stream comprises water, entrained dried metal alloy product, and any other vaporised materials. The vapour stream is cooled to condense water present. The condensed water 318 is returned to the pond 102, when the pond 102 is active. When the pond 102 is inactive, the condensed water 318 is recycled as process water, for example, for making the slurry 108.

[156] The concentration of the entrained dried metal alloy product in the vapour stream is closely monitored. If too high, the vapour is filtered (e.g. using a bag filter) to recover the entrained dried metal alloy product. The recovered dried metal alloy product is added into the dried metal alloy product that exits the rotating drum 316-1 at discharge breech 316-7.

[157] The dried metal alloy product 320 can be sold directly as an exotic metal alloy mix to specialist steel foundries. For example, where the composition of the dried metal alloy product 320 is valuable to the specialist steel foundries. Alternatively, the dried metal alloy product 320 can be subjected to further processing stages to recover particular components from the mix. For example, if the composition of the dried metal alloy product 320 is not particularly valuable.

[158] Typically, however, the dried metal alloy product 320 will be sold directly as an exotic metal alloy mix. This is because materials with high alloy content, such as the dried metal alloy product 320, offer enhanced performance properties including excellent strength and durability, and resistance to oxidation, corrosion and deforming at high temperatures or under extreme pressure. Because of these properties, exotic alloys make the best spring materials for demanding working conditions, which can be encountered across various industry sectors, including the automotive, marine and aerospace sectors as well as oil and gas extraction, thermal processing, petrochemical processing and power generation.

[159] It will be appreciated that the composition of magnetisable material in the impounded coal ash pond 102 will dictate the composition of the dried metal alloy product 320 - i.e. because those magnetisable materials in the coal ash 102 end up in the dried metal alloy product 320. This will then determine the type of exotic metal alloy that is produced. For example, when the dried metal alloy product 320 comprises significant amounts of Ni and Cu and trace amounts of Mn, it is suitable for use in the production of Monel metal. As another example, when the dried metal alloy product 320 comprises significant amounts of Fe, Cr, Ni and trace amounts of Mn, it is suitable for use in the production of 304 Stainless Steel.

[160] To further enhance the properties of the dried metal alloy product 320, Mo recovered in the acid insoluble component recovery area 600 can be added to the dried metal alloy product 320. By adding Mo to the dried metal alloy product 320, other exotic metal alloys such as Hastelloy, Inconel, Martensitic Stainless Steel, Austenitic Stainless Steel and 316 Stainless Steel can be produced.

[161] Advantageously, C^Os, CuO and NiO are all toxic compounds which are removed from the coal ash slurry in the magnetic separation stage 300 because they are magnetisable and are thus incorporated into what will be a highly valuable dried metal alloy product 320. Water-Soluble Component Recovery Area 400 - Fig. 3

[162] The slurry 304 of non-magnetisable material and substantially all the liquid which exits the drum magnetic separator 302 at the first exit, is collected and pumped to the water- soluble component recovery area 400 (Fig. 3). Because the water-soluble components dissolve from the coal ash when the coal ash is removed from the pond and mixed with water to form a pumpable slurry (see Fig. 2), the liquid component of the slurry comprises the water-soluble components, whilst the solid component of the slurry primarily comprises the non-water-soluble components.

[163] In the water-soluble component recovery area 400, the water-soluble components of the coal ash are recovered as one or more products. Some of the water-soluble components (e.g. As, B, Se) are toxic elements. As a result, typically these are recovered from the solution comprising the water-soluble components and converted into a form which is not harmful to the environment. Some of the other water-soluble components (e.g. Ba, Li, K) are not toxic, but are present in significant enough concentrations to warrant recovery as saleable by-products of the process.

[164] The slurry 304 is pumped to a dewatering centrifuge 401. The dewatering centrifuge 401 is of the same design and function as dewatering centrifuge 308. As above, the slurry 304 is passed to the rotating bowl of centrifuge 401. As the slurry is rotated, the components of the slurry separate according to density. The solid components (i.e. comprising the non- water-soluble components) are displaced further away from the axis of rotation, compared with the entrained liquor (i.e. because it is lighter than the solid). The solid components are scraped by plates to the solids outlet of the centrifuge 401 and the liquor moves counter- currently to the centrate outlet of the centrifuge 401.

[165] An ash sludge 404 (now comprising about 40-60% solids by mass) is collected from a first solids outlet of the centrifuge 401. The ash sludge 404 comprises the non- magnetisable materials (i.e. those solid materials that were not removed in the magnetic separation stage 300) and the non-water-soluble components of the coal ash. The ash sludge 404 is transported, e.g. by means of a solids conveyor, to the acid leaching stage 500.

[166] The aqueous solution comprising the water-soluble components 406 is collected from the centrate outlet of the centrifuge 401. The turbidity of the aqueous solution 406 is carefully monitored. If the turbidity of the aqueous solution is too high, the operation of the centrifuge is adjusted, and/or an additional filtration step is performed. This is because the presence of suspended solids in the feed to ion-exchange can adversely affect the operation of the ion- exchange column(s). The suspended solids can gradually plug the bed, increasing pressure losses and necessitating more frequent backwashing of the column(s).

[167] The aqueous solution 406 is pumped through a sequence of ion-exchange stages 408, 410, 420, 430, 440, 448, 456. The ion-exchange stages are carried out using one or more ion-exchange columns that are designed and operated using an ion-exchange procedure. The sequence of the ion-exchange stages is designed to sequentially remove cations in a particular order such that the resins selected in each ion-exchange stage are not adversely affected by the presence of the remaining cations in the solution.

[168] It will be appreciated that, although this specific embodiment uses exclusively ionexchange to recover each water-soluble element, other (chemical and/or physical) means may be used instead of one or more of the ion-exchange stages without departing from the spirit or scope of the process as disclosed herein.

[169] The aqueous solution 406 comprises multiple, for example up to seven, water- soluble cations: As, Ba, B, Li, K, Se, Na. Thus, there are for example up to seven ionexchange stages. Typically, the water-soluble metals are removed in the following order: As, B, Se, Ba, Li, K, Na, as per Fig. 3. It will be understood, however, that depending on the coal ash being remediated, not all these elements will be present. When a particular element is not present in a coal ash deposit, the corresponding ion-exchange stage can be omitted or isolated from the process.

[170] Each ion-exchange stage can operate in a batch or semi-continuous mode and can comprise one or more ion-exchange columns. It will be appreciated that the operating mode selected depends on a number of factors including: volume flow into the ion-exchange stage, the concentration of the element to be loaded, etc. For example, if the volume flow/concentrations are low, a batch process with only a single column may be sufficient.

However, if the volume flow/concentrations are high, a semi-continuous mode may be used with dedicated loading and elution columns.

[171] Each ion-exchange stage comprises at least one column packed with a resin. The resin is selected based on the water-soluble component of the aqueous solution 406 that is being removed in the ion-exchange stage. For example, when the component to be removed is As, a resin is selected that preferentially reacts with As in the solution. Typically, the resin comprises either protons, sodium cations or hydroxide anions that are exchanged with the component to be removed.

[172] The solution comprising the water-soluble-components is pumped into the bottom of the column packed with resin. As the solution comprising the water-soluble components is pumped through the resin loading column, the protons, sodium cations or hydroxide anions on the resin exchange with the particular water-soluble component, i.e. the particular component becomes ‘loaded’ onto the resin. Exiting the column at the top is an aqueous (barren) solution which is substantially free of the particular water-soluble component. The aqueous solution is then passed to the next column packed with resin in the sequence, i.e. for the removal of the next water-soluble component.

[173] To recover the particular water-soluble component adsorbed onto the resin, the loaded column is periodically eluted. The method of elution depends on the mode of operation, e.g. batch or semi-continuous.

[174] Fig. 3a is a schematic showing a single ion-exchange column 408-1, which operates in a batch mode. The column 408-1 comprises resin 408-2 in the form of a packed resin bed. The resin 408-2 is held in the column 408-1 by a lower support 408-3 and an upper support 408-4. Typically, the supports 408-3 and 408-4 are finely perforated to allow liquid (i.e. solution/eluant) but not resin to pass therethrough.

[175] In a batch process, the resin is first loaded by pumping the solution 406 comprising the water-soluble components from the bottom 408-5 of the column 408-1 and through the resin 408-2. The solution 406 is pumped through the resin 408-2 until a capacity of the resin 408-2 to adsorb the particular water-soluble component it preferentially reacts with is reached and/or a set period of time has elapsed. During resin loading, the solution collected from the top 408-6 of the column 408-1 is substantially free of the particular water-soluble component that reacts with the resin 408-2. The solution substantially free of the particular water-soluble component is pumped to the next ion-exchange stage in the series.

[176] Next, the resin is eluted. During elution, the solution comprising the water-soluble components is no longer pumped to the column 408-1 , for example, it may be diverted to a holding tank. Instead, a suitable eluant is pumped through the loaded column 408-1 , for a predetermined length of time. The eluant is pumped from the bottom 408-5 of the column 408-1 so it is forced to pass through the loaded resin 408-2. The eluant is selected such that, as the eluant contacts the resin, the loaded water-soluble component reacts with (i.e. exchanges with) a cation (e.g. proton or sodium) or anion in the eluant. As the loaded water- soluble component and the ion in the eluant exchange, the resin is regenerated for re-use in the loading stage and a solution comprising the particular water-soluble component is produced and collected from the top 408-6 of the column 408-1. The solution comprising the particular water-soluble component is typically a concentrated solution (concentrated eluate). Once the predetermined length of time for elution has elapsed, the column is switched back into a loading phase, e.g. the eluant is no longer pumped through the column and instead the solution comprising the water-soluble components is pumped through the column.

[177] As alluded to, the eluant is selected based on the form of the original resin used for loading the particular water-soluble component - e.g. if the resin is protonated then an acidic eluant is used, whereas if the resin contains sodium cations, a solution comprising sodium chloride is used. Typically, the eluant is pumped into the bottom of the column so that the concentrated solution comprising the water-soluble component is removed from the top of the column. Over time the resin will degrade (e.g. due to incomplete elution) and will need to be replaced.

[178] In a continuous or semi-continuous process, such as that depicted in the schematic of Fig. 3b, the ion-exchange stage comprises a dedicated resin loading column 408-7 and a dedicated elution column 408-8. The solution 406 comprising the water-soluble-components is pumped into the bottom 408-9 of the resin loading column 408-7. The loaded resin is removed (either continuously or periodically) from the bottom 408-10 of the resin loading column 408-7. The loaded resin is transferred (e.g. pumped or manually transferred) via a transfer vessel 408-21 to the top 408-11 of the elution column 408-8. A screen 408-12 allows entrained liquor to be recovered from the loaded resin. The recovered liquor is sent back to the resin loading column 408-7.

[179] A solution 408-16 substantially free of the particular water-soluble component that reacts with the resin in the loading column 408-7 is collected from an overflow 408-15 at the top of the loading column 408-7. The overflow 408-15 comprises a resin trap 408-17 which collects resin from the column that goes into the overflow. The solution 408-16 is passed to the next ion-exchange stage in the series and the collected resin is returned to the resin loading column 408-7.

[180] The resin and solution comprising the water-soluble-components move counter- currently through the resin loading column 408-7.

[181] In the elution column, a suitable eluant 408-13 is pumped through the loaded resin, regenerating the resin and producing a (concentrated) solution comprising the particular water-soluble component. The eluant 408-13 is pumped from the bottom 408-14 of the elution column 408-8 so it is forced to pass through the loaded resin. The concentrated eluate 408-18 is collected from an overflow at the top of the elution column 408-8.

Regenerated resin is removed (either continuously or periodically) from the bottom 408-14 of the elution column 408-8. That is, the resin effectively moves down the elution column 408-8 as the eluant is pumped up through the column 408-8. In this way, the loaded resin and the eluant flow counter-currently through the elution column 408-8.

[182] The regenerated resin removed from the elution column 408-8 is transferred via a transfer vessel 408-20 to the top 408-19 of the loading column 408-7. Typically, a small volume of resin (~1-5%) is lost as resin is transferred between columns and therefore a small volume of make-up resin must be added to the resin loading column.

[183] It will be appreciated that the exact operating conditions (e.g. temperature, pressure) of each ion-exchange stage depends on the resin selected. However, typically all ionexchange columns operate at elevated pressure, because there is a pressure drop within the column as the solution is pumped through it. It will be appreciated that the exact operating pressure is selected to account for such a pressure drop, i.e. so that the top of the column still has a pressure above atmospheric.

[184] Depending on the resin, other reagents may be added to help promote the ionexchange process. For example, some resins operate more effectively under alkali conditions, so a suitable reagent is added to these ion-exchange stages to increase the pH. Conversely, some resins operate more effectively under acidic conditions, so a suitable reagent is added to these ion-exchange stages to decrease the pH.

[185] Referring again to Fig. 3, the aqueous solution 406 is pumped through seven ionexchange stages in series 408, 412, 420, 430, 440, 448, 456. Typically, each of these ionexchange stages operates as a batch process, i.e. as described with reference to Fig. 3a above.

[186] The aqueous solution 406 is pumped to the column of the first ion-exchange stage 408. The column of the first ion-exchange stage 408 is loaded with a resin that preferentially reacts with As. The resin is typically a strong-base anion exchange resin, because As is in the form of HAsC ² ' in solution.

[187] As above, as the aqueous solution 406 is pumped through the column, the As reacts with the strong-base anion exchange resin and loads onto the resin. A solution 410 substantially free of As is pumped from the top of the column to the second ion-exchange stage 412.

[188] The resin (now loaded with As) is eluted. As above (for a batch process), the aqueous solution 406 stops being pumped into the column once the resin is sufficiently loaded with As (or after a predetermined time has elapsed) and instead a suitable eluant is pumped through the column. A suitable eluant for regenerating a strong-base anion exchange resin is NaOH at a concentration of about 4%. The As exchanges with hydroxide anions in the eluant, regenerating the resin and producing a concentrated solution comprising arsenic.

[189] The concentrated eluate (i.e. a concentrated solution of As) 414 is pumped from the top of the column to a tank 415 for storage. The tank 415 is periodically emptied of the concentrated solution of As (i.e. as it fills up), with the concentrated solution of As 416 optionally then being used for (e.g. added to) the manufacture of a cementitious product. Once the resin is regenerated (or after a predetermined time), the pump pumping the eluant is stopped and aqueous solution 406 can again be pumped into the column 408.

[190] The solution 410 substantially free of As is pumped to the column of the second ionexchange stage 412. The column of the second ion-exchange 412 is loaded with a resin that preferentially reacts with B, for example AmberLite PWA10 resin. As above, as the solution 410 substantially free of As is pumped through the column, B in the solution 410 reacts with the resin, thereby loading onto the resin. A solution 418 substantially free of B (and As) is pumped from the top of the column to the third ion-exchange stage 420.

[191] The resin (now loaded with B) is eluted. As above, the solution 410 stops being pumped into the column once the resin is sufficiently loaded with B (or after a predetermined time has elapsed). A suitable eluant, such as sodium hydroxide, is then pumped through the column, regenerating the resin. A concentrated eluate (i.e. a concentrated solution of B) 422 is pumped from the top of the column to a tank 424 for storage. The tank 424 is periodically emptied of the concentrated solution of B (i.e. as it fills up), with the concentrated solution of B 426 added to the manufacture of a cementitious product.

[192] Once the resin is regenerated (or after a predetermined time), the pump pumping the eluant is stopped. Typically, when AmberLite PWA10 resin is used, the second ionexchange stage 412 comprises an additional conversion step after elution. This occurs because the regenerated resin from elution is not regenerated to its original form and therefore does not comprise the functional groups (i.e. the ions on the resin that react with the B to cause the ion-exchange) required to react with B during loading. For example, the resin may be regenerated with protons although sodium cations are preferred when loading the resin, because the backwards exchange of B for sodium occurs slowly compared to the exchange of B for protons. The conversion step comprises pumping a suitable solution through the regenerated resin that converts the functional groups back to their original form. Once conversion is complete, the solution 410 is again pumped to the column to load the resin. [193] The solution 418 substantially free of B is pumped to the column of the third ionexchange stage 420. The column of the third ion-exchange stage 420 is loaded with a resin that preferentially reacts with Se. Typically, Se ions can be selectively removed using a metal hydroxide loaded ion-exchange resin or with chelating resins loaded with oxyanions. As above, as the solution 418 is pumped through the column, Se in the solution 418 reacts with the resin and loads onto the resin. A solution 428 substantially free of Se (and As, B) is pumped from the top of the column to the fourth ion-exchange stage 430.

[194] The resin (now loaded with Se) is eluted. As above, the solution 418 stops being pumped into the column and a suitable eluant, such as sodium hydroxide, is then pumped through the column, regenerating the resin. A concentrated (Se) eluate 432 is pumped from the top of the column to a tank 434 for storage. The tank 434 is periodically emptied of the concentrated solution of Se (i.e. as it fills up), with the concentrated solution of Se 436 added to the manufacture of a cementitious product.

[195] Once the resin is regenerated, the pump pumping the eluant is stopped, and the column is again used for loading, i.e. by pumping solution 418 into the regenerated column.

[196] Advantageously, the water-soluble component recovery area 400 allows the toxic elements As, B and Se to be extracted and recovered from the coal ash. The recovered toxic elements can be stabilised by incorporating them into a cementitious product, i.e. by using solutions from tanks 415, 424 and 434 as components of said cementitious product. The cementitious product can comprise, for example, the structural light weight aggregate described in AU 2021902692. When the solutions from tanks 415, 424 and 434 are used as a component of the structural light weight aggregate, the solutions 416, 426 and 436 are pumped to the dryer prior to the particle separation stage of the process of AU 2021902692.

[197] However, depending on the concentration of As, B and Se, the solutions from tanks 416, 426 and 436 may be suitable for use in the manufacture of other cementitious products. For example, the solutions 416, 426, 436 can be used in the manufacture of cement roof tiles, cement blocks and/or concrete panels.

[198] The solution 428 substantially free of Se is pumped to the column of the fourth ionexchange stage 430. The column of the fourth ion-exchange stage 430 is loaded with a resin that preferentially reacts with Ba. Suitable resins include AmberLite HPR1100 strong acid cation exchange resin (loaded with sodium cations) or AmberLite IRC83 H weak acid cation exchange resin (which may be loaded with either protons or sodium cations). As above, as the solution 428 is pumped through the column, Ba in the solution 428 reacts with the resin and loads onto the resin. A solution 438 substantially free of Ba (and As, B, Se) is pumped from the top of the column to the fifth ion-exchange stage 440.

[199] The resin (now loaded with Ba) is eluted. As above, the solution 428 stops being pumped into the column and a suitable eluant, such as hydrochloric acid, is then pumped through the column, regenerating the resin. A concentrated (Ba) eluate 442 is pumped from the top of the column to a tank 444 for storage. The tank 444 is periodically emptied of the concentrated solution of Ba (i.e. as it fills up), with the concentrated solution of Ba sold as a valuable by-product of the process 10. Optionally, the concentrated solution of Ba from tank 444 can be dried prior to sale, i.e. to decrease the moisture content of the final product.

[200] Once the resin is regenerated, the pump pumping the eluant is stopped, and the column is again used for loading, i.e. by pumping solution 428 into the column.

[201] The solution 438 substantially free of Ba is pumped to the column of the fifth ionexchange stage 440. The column of the fifth ion-exchange stage 440 is loaded with a resin that preferentially reacts with Li. Suitable resins include strong acid cation exchangers, such as AmberSep G26 H resin. As above, as the solution 438 is pumped through the column, Li in the solution 438 reacts with the resin and loads onto the resin. A solution 446 substantially free of Li (and As, B, Se, Ba) is pumped from the top of the column to the fifth ion-exchange stage 448.

[202] The resin (now loaded with Li) is eluted. As above, the solution 438 stops being pumped into the column and a suitable eluant, such as hydrochloric acid, is then pumped through the column, regenerating the resin. A concentrated (Li) eluate 450 is pumped from the top of the column to a tank 452 for storage. The tank 452 is periodically emptied of the concentrated solution of Li (i.e. as it fills up), with the concentrated solution of Li sold as a valuable by-product of the process 10. Optionally, the concentrated solution of Li from tank 452 can be dried prior to sale, i.e. to decrease the moisture content of the final product.

[203] Once the resin is regenerated, the pump pumping the eluant is stopped, and the column is again used for loading, i.e. by pumping solution 438 into the (regenerated) column.

[204] The solution 446 substantially free of Li is pumped to the column of the sixth ionexchange stage 448. The column of the sixth ion-exchange stage 448 is loaded with a resin that preferentially reacts with K. Suitable resins include AmerLite FPC88 UPS Strong Acid Cation Resin which converts salts to acids. Alternatively, Chromatographic Separation Resin Size of 350 pm can be used to separate K from the remaining Na in solution 446. As above, as the solution 446 is pumped through the column, K in the solution 446 reacts with the resin and loads onto the resin. A solution 454 substantially free of K (and As, B, Se, Ba, Li) is pumped from the top of the column to the fifth ion-exchange stage 456.

[205] The resin (now loaded with K) is eluted. As above, the solution 446 stops being pumped into the column and a suitable eluant, such as hydrochloric acid, is then pumped through the column, regenerating the resin. A concentrated (K) eluate 462 is pumped from the top of the column to a tank 464 for storage. The tank 464 is periodically emptied of the concentrated solution of K (i.e. as it fills up), with the concentrated solution of K sold as a valuable by-product of the process 10. Optionally, the concentrated solution of K from tank 464 can be dried prior to sale, i.e. to decrease the moisture content of the final product.

[206] Once the resin is regenerated, the pump pumping the eluant is stopped, and the column is again used for loading, i.e. by pumping solution 446 into the (regenerated) column.

[207] The solution 454 substantially free of K is pumped to the column of the seventh ionexchange stage 456. The solution now primarily comprises Na only, i.e. because the As, Be, Se, Ba, Li, K have been extracted via ion-exchange stages 408, 412, 420, 430, 440, 448 respectively. Suitable resins for extracting the Na include AmberLite FPC88 UPS Strong Acid Cation Resin. As above, as the solution 454 is pumped through the column, Na in the solution 454 reacts with the resin and loads onto the resin. A solution 458 substantially free of Na (and As, B, Se, Ba, Li, K) is recovered from the top of the column. Because the solution 458 is substantially free of all the water-soluble components, it is suitable to be recycled back to the pond 102, when the pond is active. When the pond 102 is inactive, the solution 458 can be recycled for use as process water within the process 10.

[208] The resin (now loaded with Na) is eluted. As above, the solution 454 stops being pumped into the column and a suitable eluant, such as hydrochloric acid, is then pumped through the column, regenerating the resin. A concentrated (Na) eluate 460 is pumped from the top of the column and is sent to the chlor-alkali plant for re-use.

[209] Once the resin is regenerated, the pump pumping the eluant is stopped, and the column is again used for loading, i.e. by pumping solution 454 into the (regenerated) column.

[210] It will be appreciated that other separation techniques for preferentially separating and recovering each of the water-soluble metals can be employed, including e.g. a step-wise pH-regulated precipitation regime.

Acid Leaching Area 500 - Fig. 4 [211] Referring now to Fig. 4, the ash sludge 404 comprising the non-magnetisable materials and the non-water-soluble components of the coal ash (i.e. the remaining coal ash residue) is transported by means of a solids conveyor to the acid leaching stage 500. In the acid leaching stage 500, a hydrometallurgical leaching process is performed to leach acidsoluble components from the coal ash sludge 404.

[212] The ash sludge 404 is added to a digestion vessel 503. Hydrochloric acid 504 is also added thereto. As the hydrochloric acid 504 is added to the digestion vessel 503, the hydrochloric acid 504 and the coal ash sludge 404 form a slurry. Typically, the hydrochloric acid 504 has a concentration of about 10% by mass.

[213] The digestion vessel 503 comprises an agitator which is operated to ensure homogeneity within the vessel by continuously mixing the contents of the vessel. The agitator is powered by a variable speed motor such that the speed of the agitator can be adjusted as required.

[214] The digestion vessel 503 is operated at a temperature of about 90 °C and at atmospheric pressure. The digestion vessel 503 is typically jacketed, with steam flowing through the jacket, so as to indirectly heat the contents of the vessel. As the steam flows through the jacket, the steam is condensed, and energy is transferred to the contents of the vessel. The temperature of 90 °C is maintained by controlling the flow of steam through the jacket.

[215] The reaction temperature and the residence time are selected based on the reaction kinetics so as to maximise throughput and minimise cost. The volume of the digestion vessel 503 is selected based on the residence time required and the volumetric flow of slurry exiting the reactor (i.e. reactor volume equals volumetric flow divided by residence time, plus any headspace requirement within the vessel).

[216] The digestion vessel 503 can comprise several digestion vessels in series. The liquid overflow from the first digestion vessel is fed into a feed inlet of the next digestion vessel, and so on. The leaching process is operated continuously whereby slurry continuously flows (or is pumped) from one tank to the next. The number and size of the digestion vessels is selected based on the residence time required for the leaching reactions to achieve the required extent of metal leaching.

[217] Alternatively, the leaching process can be operated as a batch process. When the leaching process is operated as a batch process, a predetermined amount of the ash sludge 404 is loaded into the digestion vessel 503. A predetermined mass of 10% (by mass) hydrochloric acid 504 is next added into the digestion vessel 503. The digestion vessel is heated to about 90 °C for a predetermined period of time. The predetermined period of time is selected based on the reaction kinetics so as to maximise leaching of the acid-soluble components from the coal ash residue 404. After the predetermined time has elapsed, the contents of the digestion vessel 503 are removed.

[218] As the ash sludge 404 and hydrochloric acid 504 are mixed in the digestion vessel 503 and are subjected to the elevated temperature (as part of either a continuous or batch process), many of the metal oxides (MO) in the ash sludge 404 will react with the hot dilute hydrochloric acid to give metal chlorides (MCI) via (Reaction 1).

MO + n HCI(aq) M ⁿ⁺ + n Ch + H ₂ O (R1)

[219] Some of the metal chlorides are soluble in the hydrochloric acid (e.g. Sb, Be, Cd, Zn, Sn) and form a solution comprising the acid-soluble components of the ash sludge 404. However, some of the metal chlorides are insoluble in the hydrochloric acid (e.g. Hg, Pb) and precipitate from solution. The remaining solid residue after leaching thus comprises the remaining non-acid-soluble components (i.e. those oxides and other species that did not react with the hydrochloric acid to form chlorides) and precipitated chlorides (i.e. those metal chlorides that are insoluble in an acidic environment).

[220] Hydrochloric acid 504 is added to the digestion vessel 503 so as to maintain an acidic pH within the digestion vessel 503. The pH within the vessel 503 can be measured, for example, with a pH probe. Hydrochloric acid 504 can be added until a predetermined pH is achieved. Typically, the (total) mass of hydrochloric acid 504 added to the digestion vessel 503 is equivalent to the mass of solids in the ash sludge 404 added to the digestion vessel 503.

[221] An acidic slurry 506 comprising the solution comprising the acid-soluble components of the coal ash residue and the solid comprising the non-acid-soluble components and precipitated chlorides is pumped from the digestion vessel 503 to a dewatering centrifuge 508. The dewatering centrifuge 508 is of the same type and function as dewatering centrifuge 308 and made of a material that is resistant to the corrosive nature of hydrochloric acid.

[222] As above, in the dewatering centrifuge 508, the acidic solution 502 comprising the acid-soluble components of the ash sludge is separated from the ash sludge 505 comprising the non-acid-soluble components and precipitated chlorides. The acidic solution 502 is collected from the centrate outlet of dewatering centrifuge 508 and is pumped to the acidsoluble recovery stage 700 (i.e. Fig. 7). The ash sludge 505 is collected from the solids outlet of the dewatering centrifuge 508 and is transported, e.g. using a solids conveyor, to the non-acid-soluble recovery area 600 (i.e. Figs. 5 and 6). The ash sludge 505 typically comprises about 40-60% solids by mass.

[223] Hydrochloric acid is the preferred acid for the digestion because it enables the conversion of (at least) Hg and Pb oxides to chlorides, which precipitate as above. Hydrochloric acid is also non-oxidising, so does not degrade the ion-exchange resins present in the acid-soluble recovery stage 700.

Non-Acid-Soluble Component Recovery Area 600 - Figs. 5 and 6

[224] The ash sludge 505 is transported to the dissolution vessel 602 of the non-acid- soluble component recovery area 600. Pond coal ash comprises Hg and Pb, both of which are toxic and not soluble in acid. These components must be extracted from the coal ash and stabilised in forms which will not be harmful to the environment to ensure the alumina and silica products are free from toxic contaminants. Pond coal ash also comprises other components which are not soluble in acid and, whilst not toxic, are present in significant enough quantities to warrant recovery of these components as saleable by-products. These other non-acid-soluble components typically comprise Ge, Au, Ag, Mo, Ti, Si and Zr. a. The non-acid-soluble component recovery area 600 is comprised of (up to) six treatment stages in which one or more of the non-acid-soluble components are recovered as a respective product. As shown in Figs. 5 and 6, the non-acid-soluble component recovery area 600 typically comprises the following six stages: b. Hg and Pb separation and recovery; c. Precious metals separation and recovery; d. Ti separation and recovery; e. Mo separation and recovery; f. Ge separation and recovery; and g. Zr and Si separation and recovery.

[225] For clarity, in Figs. 5 and 6, the bolded lines indicate the main process lines, that is the direction the silica moves within the process.

[226] It will be appreciated that the order and number of the stages in the non-acid-soluble component recovery area 600 can vary, depending on the composition of the impounded coal ash being remediated. For example, if the impounded coal ash does not comprise any precious metals, then stage (ii) is omitted or isolated. Alternative, if the impounded coal ash does not comprise any Mo, then stage (iv) is omitted or isolated. Of course, if the impounded coal ash does not comprise any of the non-acid-soluble components Hg, Pb, Ge, Au, Ag, Mo, Ti, Si and Zr, then the non-acid-soluble component recovery area 600 can be omitted or isolated from the process altogether, although noting it would be rare/surprising for silica not to be present in coal ash.

[227] Each of the stages (i) to (vi) may be carried out under a range of conditions, using a range of different approaches. A specific embodiment of each stage is described in detail below. However, other approaches may be adopted by a person of ordinary skill in the art to achieve the desired result.

(i) Hg and Pb recovery - Fig. 5

[228] The ash sludge 505 is added to the dissolution vessel 602. The contents of the dissolution vessel 602 are maintained at a temperature of about 90 °C, for example by indirectly heating the dissolution vessel 602 by means of a steam-jacket, as described above. As set forth above, the ash sludge 505 comprises the non-acid-soluble components of the coal ash (e.g. alumina and silica) and precipitated chlorides. Both Hg and Pb are present in ash sludge 505 as precipitated chlorides (i.e. HgCh and PbCl2 solids respectively) which are soluble in hot, neutral solutions. The ash sludge 505 is acidic because of the acidic liquid entrained therewith (i.e. the remaining 40-60% by mass of the ash sludge 505 that is not solid).

[229] Sodium hydroxide 606 at a concentration of about 10% by mass is added to the dissolution vessel 602. As the sodium hydroxide 606 is added, hydrochloric acid present in the dewatered solid 505 is neutralised. Sodium hydroxide 606 is incrementally added until the pH is around 7, i.e. neutral. This is because, by having an approximately neutral pH within the dissolution vessel 602, the Hg and Pb chlorides can be caused to dissolve from the dewatered solid 505, i.e. because they are soluble in hot, neutral solutions. However, any remaining chlorides and the non-acid-soluble components remain as a solid residue. The exact target pH is selected so as to maximise the dissolution of the Hg and Pb chlorides.

[230] Hot water 604 at a temperature of about 90 °C (i.e. the operating temperature of the dissolution vessel 602) is added to help promote dissolution of Hg and Pb chlorides and to increase the liquids fraction in the dissolution vessel 602. This allows the dissolved Hg and Pb chlorides to form a solution comprising the Hg and Pb chlorides. [231] The dissolution vessel 602 typically comprises some form of agitation, such as an agitator attached to a variable speed motor. Agitation of the contents of the dissolution vessel 602 ensures homogeneity within the dissolution vessel 602 and helps maximise the efficiency of the dissolution reaction. The size of the dissolution vessel 602 is selected to produce sufficient residence time for the dissolution reactions to achieve the maximum extent, thereby maximising recovery of the Hg and Pb into solution.

[232] As above, depending on the volume of dewatered solid 505 and the residence time required, the dissolution can comprise a series of such dissolution vessels in series. Exiting the dissolution vessel 602 (or the final dissolution vessel in a series of dissolution vessels), is a slurry 608. The solid component of the slurry 608 comprises a solid residue comprising the non-acid-soluble components and the remaining chlorides. The liquid component of the slurry 608 is a solution comprising the hot water-soluble chlorides (i.e. Hg and Pb).

[233] The slurry 608 is passed, e.g. by a slurry pump or by gravity, to a centrifuge bank 610. The centrifuge bank 610 comprises one or more centrifuges which operate in parallel. The slurry 608 is divided between the one or more centrifuges in the centrifuge bank 610.

For example, the slurry 608 is divided based on which centrifuge(s) are operating and/or the capacity of the centrifuge(s) within the centrifuge bank 610.

[234] As the slurry 608 passes into the centrifuge bank 610, a sludge 612 comprising the non-acid-soluble components and the remaining chlorides is separated from a solution 614 comprising the water-soluble chlorides. The liquor content of the sludge 612 is typically around 50% by mass. To increase the recovery of the water-soluble chlorides, the centrifuge bank 610 includes an optional wash stage.

[235] In the wash stage, the sludge 612 is washed with fresh liquor (e.g. fresh water or recycled liquor from within the process) which comprises minimal impurities. As the sludge 612 is washed with fresh liquor comprising minimal impurities, the solution comprising the water-soluble chlorides entrained within the solid residue, i.e. as interstitial liquor, is displaced by the fresh liquor comprising minimal impurities. The displaced solution comprising the water-soluble chlorides is recovered and added to the solution 614 comprising the water-soluble chlorides.

[236] The sludge 612 is transported, e.g. by a solids conveyor, to stage (ii) precious metals separation and recovery.

[237] The solution 614 comprising the water-soluble chlorides is pumped to the column of a Pb and Hg ion-exchange stage 616. The ion-exchange stage 616 is typically configured as a batch process and is operated and configured similarly to the ion-exchange stages in the water-soluble component recovery area 400.

[238] The column of the Pb and Hg ion-exchange stage 616 is loaded with a resin that preferentially reacts with both Pb and Hg. As above, as the solution 614 is pumped through the column, Pb and Hg in the solution 614 reacts with the resin and loads onto the resin. A suitable resin for loading both Pb and Hg is AmberSep GT75 chelating resin. A solution 620 substantially free of Pb and Hg is recovered from the top of the column and can be reused in the process, e.g. as process water and/or returned to the pond 102.

[239] The resin (now loaded with Pb and Hg) is eluted. As above, the solution 614 stops being pumped into the column and a suitable eluant, such as concentrated hydrochloric acid, is instead pumped through the column, regenerating the resin. A concentrated (Pb and Hg) eluate 624 is pumped from the top of the column to a Pb and Hg recovery stage 622.

[240] Once the resin is regenerated, the pump pumping the eluant is stopped, and the column is again used for loading, i.e. by pumping solution 614 into the (regenerated) column.

[241] In the Pb and Hg recovery stage 622, the Pb and Hg in the concentrated eluate 624 are recovered as chlorides 647. First, sodium hydroxide 632 is added to the concentrated eluate 624 to neutralise any acid remaining in solution. By neutralising the acid, the concentrated eluate 624 becomes neutral and the Pb and Hg are caused to precipitate therefrom as chlorides.

[242] The precipitate 647 comprising the Pb and Hg chlorides is separated from a remaining solution 649 by filtration. The separated precipitate 647 can optionally be dried, e.g. in an indirectly- heated rotary dryer and can be incorporated as a component of a cementitious product. The Pb and Hg precipitate 647 can also be used as a component of a cementitious product. For example, the cementitious product can comprise the structural light weight aggregate described in AU 2021902692. When the Pb and Hg precipitate is used as a component of the structural light weight aggregate, the concentrated eluate 624 (i.e. comprising the Hg and Pb in solution) is typically pumped directly to a dryer prior to the particle separation stage of the process of AU 2021902692. As well as being used in a structural light weight aggregate, the precipitate 647 may be suitable for use in a range of other cementitious products including roof tiles, cement blocks, concrete panels, etc., provided any excess acid is first neutralised with say NaOH solution.

[243] The separated neutralised liquor 649 is essentially NaCI solution and can be reused in the chlor-alkali plant or as an IX regenerative solution. [244] It will be appreciated that, because AmberSep GT75 chelating resin is selected for use in the ion-exchange stage 616, both Pb and Hg are extracted in a single ion-exchange stage. However, two ion-exchange stages could instead be used with resins selected that preferentially react with Pb or Hg. That is, a separate Pb concentrated eluate and a separate Hg concentrated eluate would instead be produced. This is beneficial where there are separate end-uses for each product.

[245] It will be appreciated that other separation techniques for preferentially separating and recovering Pb and Hg can be employed. For example, the solution 614 can be reacidified with hydrochloric acid so as to cause the Pb and Hg chlorides to precipitate therefrom. The precipitate comprising Pb and Hg chlorides is then separated from a residual acidic solution. As above, the separated Pb and Hg chloride precipitate is suitable for use as a component of a cementitious product. The separated residual acidic solution is recovered and pumped to an HCI recovery tank.

(ii) Precious metals separation and recovery - Fig. 5

[246] The sludge 612, now substantially free of Pb and Hg, is transported to a digestion vessel 626 in stage (ii) of the non-acid-soluble recovery area 600. In the digestion vessel 626, sodium thiosulphate 628 is added to the sludge 612. The sodium thiosulphate 628 typically has a concentration of about 10% sodium thiosulphate by mass. The addition of sodium thiosulphate 628 to the sludge 612 promotes the leaching of Au and Ag (when present) from the sludge into solution. The amount of sodium thiosulphate 628 added to the sludge 612 is therefore selected based on the composition of Au and Ag within the sludge 612, i.e. to ensure sufficient sodium thiosulphate is present to leach substantially all of the Au and Ag present in the sludge.

[247] Depending on the solids fraction in the digestion vessel 626, water 630 is optionally added to the digestion vessel 626. For example, when the solids fraction in the digestion vessel 626 is above a predetermined solids fraction, water 630 is incrementally added to the digestion vessel 626 until the predetermined solids fraction is achieved. However, if the solids fraction in the digestion vessel 626 is below the predetermined solids fraction, then additional water is not required. The predetermined solids fraction is typically between about 30-40% by mass solids.

[248] The digestion vessel 626 typically comprises some form of agitation, such as an agitator attached to a variable speed motor. Agitation of the contents of the digestion vessel 626 ensures homogeneity within the digestion vessel 626 and helps maximise the efficiency of the leaching reaction. The size of digestion vessel 626 is selected to produce sufficient residence time for the leaching of Au and Ag to achieve the required extent of leaching.

[249] The leaching of Au and Ag with sodium thiosulphate is more effective under slightly alkali conditions (i.e. within a pH range of about 8 to about 10). If required, sodium hydroxide is optionally added to the digestion vessel 626, i.e. to maintain a pH within the range of about 8 to about 10 within the digestion vessel 626. This can maximise the recovery of Au and Ag from the solid residue. However, typically, the addition of sodium hydroxide into the digestion vessel 626 is not required, because the sodium hydroxide 606 added to the dissolution vessel 602 is sufficient to ensure the residual sludge 612 is slightly alkali.

[250] Exiting the digestion vessel 626 is a slurry 634. The solid component of the slurry 634 comprises a residual sludge substantially free of precious metals. The liquid component of the slurry 634 comprises a solution comprising the precious metals.

[251] The slurry 634 is passed, e.g. by a slurry pump or by gravity, to a centrifuge bank 636. As above, the centrifuge bank 636 comprises one or more centrifuges which operate in parallel. The slurry 634 is divided between the one or more centrifuges in the centrifuge bank 636. For example, the slurry 634 is divided based on which centrifuge(s) are operating and/or the capacity of the centrifuge(s) within the centrifuge bank 636.

[252] As the slurry 634 passes into the centrifuge bank 636, the residual sludge 638 substantially free of precious metals is separated from the solution 640 comprising the precious metals. The liquor content of the residual sludge 638 is typically about 50% solids by mass. To increase the recovery of the precious metals, the centrifuge bank 636 includes an optional wash stage.

[253] As above, in the wash stage, the residual sludge 638 is washed with fresh liquor (e.g. fresh water or recycled liquor from within the process) which comprises minimal impurities. As the residual sludge 638 is washed with fresh liquor comprising minimal impurities, the solution comprising the precious metals entrained within the solid residue, i.e. as interstitial liquor, is displaced by the fresh liquor comprising minimal impurities. The displaced solution comprising the precious metals is recovered and added to the solution 640 comprising the precious metals.

[254] The solution 640 comprising the precious metals is transported, e.g. by pump, to a gold electrolysis chamber 642. In the gold electrolysis chamber 642, the solution comprising the precious metals 640 is subjected to a current under which gold preferentially deposits onto carbon cathode(s) placed in the solution. A layer of deposited gold builds up on the surface of the carbon cathode(s) as the gold is preferentially deposited onto the surface. [255] To recover the gold, the carbon cathode(s) are periodically removed from the gold electrolysis chamber 642. The carbon cathode(s) with the deposited gold can be sold directly as a product. Alternatively, the deposited gold can be removed from the surface of the carbon cathode(s), such as by scraping the carbon cathode(s). The scraped gold is sold as a valuable gold by-product 644 of the process and the carbon cathode(s) are recovered for reuse.

[256] As the gold is deposited onto the surface of the carbon cathode(s), the solution comprising precious metals in the gold electrolysis chamber 642 becomes depleted of gold. The gold-depleted solution 646 is pumped to a silver electrolysis chamber 648. In the silver electrolysis chamber 648, the gold-depleted solution is subjected to a current under which silver preferentially deposits onto carbon cathode(s) placed in the solution. A layer of deposited silver builds up on the surface of the carbon cathode(s) as the silver is preferentially deposited onto the surface.

[257] To recover the silver, the carbon cathode(s) are periodically removed from the silver electrolysis chamber 648. The carbon cathode(s) with the deposited silver can be sold directly as a product. Alternatively, the deposited silver can be removed from the surface of the carbon cathode(s), such as by scraping the carbon cathode(s). The scraped silver is sold as a valuable silver by-product 650 of the process.

[258] As the silver is deposited onto the surface of the carbon cathode(s), the gold- depleted solution in the silver electrolysis chamber 648 also becomes depleted of silver. The solution 652 remaining after silver electrolysis is therefore substantially free of precious metals. The solution 652 is pumped from the silver electrolysis chamber 648 and is collected for reuse within the process. Typically, the solution 652 is used to generate the solution 628 comprising sodium thiosulphate that is added to the digestion vessel 626 - i.e. the liquor operates in a closed-loop because the solution 652 recovered from the silver electrolysis chamber 648 is re-used to produce the sodium thiosulphate solution 628.

[259] As the solution 652 is continuously recycled, the concentration of contaminants within the solution 652 is continuously monitored. This is because contaminants can build-up within the solution 652 as it is recycled. When the concentration of contaminants reaches a predetermined threshold, some of the solution 652 is bled from the circuit. Depending on the composition of the bleed stream, it may require further processing (i.e. to remove the contaminants) before being disposed (e.g. back into the pond 102).

(Hi) Ti separation and recovery - Fig. 5 [260] The residual sludge 638 substantially free of the precious metals is transported from the centrifuge bank 636 to a nitric acid digester 654 in stage (iii) of the non-acid-soluble recovery area 600.

[261] In the nitric acid digester 654, concentrated nitric acid 656 is added to the residual sludge 638. The concentrated nitric acid 656 typically has a concentration of 68% by mass nitric acid (i.e. approximately the azeotropic concentration).

[262] The conditions of the nitric acid digester 654 are controlled so as to promote the dissolution of nitric acid-soluble components from the residual sludge. Typically, the nitric acid-soluble components will comprise Ti. However, depending on the composition of the impounded coal ash to the process, other nitric acid-soluble components may also be present.

[263] Ti present in the residual sludge 638 is in the form of titanium dioxide (TiC>2). As the nitric acid 656 contacts the titanium dioxide, the titanium dioxide dissolves to form titanium nitrate (Ti(NO ₃ )4), according to (Reaction 2):

TiO ₂ + 4 HNO ₃ Ti(NO ₃ ) ₄ + 2 H ₂ O (R2)

[264] Titanium nitrate is soluble in concentrated nitric acid, so will remain in solution. Other oxides that may be present in the residual sludge 638 that are soluble in nitric acid will likewise react with the concentrated nitric acid forming metal nitrates. Components in the residual sludge that do not react with nitric acid will remain in the sludge in solid form.

[265] Notably, nitric acid is only used as an acid for the purposes of separating and recovering Ti, i.e. in all other process stages requiring an acid, hydrochloric acid is preferred. This is because titanium dioxide is not soluble in hydrochloric acid.

[266] The amount of nitric acid 656 added to the nitric acid digester 654 is controlled so as to maximise the fraction of Ti (and other nitric acid-soluble components) recovered from the residual sludge. For example, a set volume/mass of nitric acid 656 can be added to the nitric acid digester 654 based on the mass of residual sludge 638 and the concentration of Ti (and other nitric acid-soluble components) within the sludge, so as to ensure an excess of nitric acid within the nitric acid digester 654. Alternatively, nitric acid 656 can be added to the nitric acid digester 654 so as to achieve a set pH within the nitric acid digester 654. The set pH is selected such that an excess of nitric acid is present within the nitric acid digester 654. The (total) mass of (concentrated) nitric acid 656 added to the digester 654 is typically equal to the mass of solids in the residual sludge 638. [267] The nitric acid digester 654 typically comprises some form of agitation, such as an agitator attached to a variable speed motor. Agitation of the contents of the nitric acid digester 654 ensures homogeneity within the nitric acid digester 654 and helps maximise the efficiency of the reaction between the titanium oxide in the residual sludge and the concentrated nitric acid. The size of the nitric acid digester 654 is selected to produce sufficient residence time for the reaction between the titanium dioxide and concentrate nitric acid to achieve the maximum extent, thereby maximising recovery of Ti into solution.

[268] Exiting the nitric acid digester 654 is a slurry 656. The solid component of the slurry 656 comprises a residual sludge which is substantially free of the nitric acid-soluble components (e.g. Ti). The liquid component of the slurry 656 is a solution comprising the nitric acid-soluble components (e.g. Ti).

[269] The slurry 656 is passed, e.g. by a slurry pump or by gravity, to a centrifuge bank 658. As above, the centrifuge bank 658 comprises one or more centrifuges which operate in parallel. The slurry 656 is divided between the one or more centrifuges in the centrifuge bank 658.

[270] As the slurry 656 passes into the centrifuge bank 658, a residual sludge 660 substantially free of nitric acid-soluble components is separated from a solution 662 comprising the nitric acid-soluble components. The residual sludge 660 comprises about 40- 60% by mass solids. To increase the recovery of the nitric acid-soluble components (which are present in the interstitial solution of the residual sludge 660), the centrifuge bank 658 can include an optional wash stage.

[271] In the wash stage, the residual sludge 660 is washed with fresh liquor (e.g. fresh water or recycled liquor from within the process) which comprises minimal impurities. As above, as the residual sludge 660 is washed with fresh liquor comprising minimal impurities, the solution comprising the nitric acid-soluble components entrained within the solid residue is displaced by the fresh liquor. The displaced solution comprising the nitric acid-soluble components is recovered and added to the solution 662 comprising the nitric acid-soluble components.

[272] The residual sludge 660 is transported, e.g. by solids conveyor, to the next stage (iv) of the non-acid-soluble recovery area 600 (see Fig. 6).

[273] The solution 662 comprising the nitric acid-soluble components is transported, e.g. by liquor pump, to a precipitation tank 664. [274] The solution 662 comprising the nitric acid-soluble components is acidic, because of the addition of nitic acid to the nitric acid digester 654. In the precipitation tank 664, sodium hydroxide 666 is added to the solution 662 comprising the nitric acid-soluble components. The sodium hydroxide 666 has a concentration of 45% by mass (i.e. approximately 17 M) and is added until all the free nitric acid is neutralised.

[275] As the sodium hydroxide 666 is added to the precipitation tank 664, the sodium hydroxide 666 reacts with the titanium nitrate in solution thereby forming titanium hydroxide (Ti(OH)4), according to (Reaction 3):

Ti(NO ₃ ) ₄ + 4 NaOH Ti(OH) ₄ + 4 NaNO ₃ (R3)

[276] Titanium hydroxide is insoluble under neutral conditions/alkali and so the Ti is caused to reprecipitate from the solution, as the insoluble hydroxide. Other nitric acid-soluble components that are present in the solution 662 will similarly react with sodium hydroxide.

[277] Sufficient sodium hydroxide 666 is added to ensure the nitric acid is neutralised, the titanium is converted to a hydroxide and the titanium hydroxide is reprecipitated. Typically, the mass of sodium hydroxide 666 added to the precipitation tank 664 is about 1.3 times the mass of concentrated nitric acid 656 added to the nitric acid digester 654.

[278] The precipitation tank 664 typically comprises some form of agitation, such as an agitator attached to a variable speed motor. Agitation of the contents of the precipitation tank 664 ensures homogeneity within the precipitation tank 664 and helps maximise the efficiency of the reaction between the titanium oxide in the residual sludge and the concentrated nitric acid. The size of the precipitation tank 664 is selected to produce sufficient residence time for the reaction between the titanium nitrate and sodium hydroxide to achieve the maximum extent, thereby maximising recovery of Ti into the reprecipitated solid.

[279] Exiting the precipitation tank 664 is a slurry 668. The solid component of the slurry 668 comprises a reprecipitated solid comprising the nitric acid-soluble components (e.g. Ti). The liquid component of the slurry 668 comprises a neutralised solution, which is substantially free of the nitric acid-soluble components (e.g. Ti).

[280] The slurry 668 is pumped by a slurry pump to a filter 670. For example, the filter 670 can be a belt filter or a filter press. The type of filter 670 is selected depending on the volume of the slurry 668, the solids fraction of the slurry and the properties of the reprecipitated solid in the slurry.

[281] As the slurry 668 is passed through the filter 670, the solid component of the slurry comprising the reprecipitated solid is separated as a filter cake from the liquid component of the slurry comprising the neutralised solution. When the filter 670 is a belt filter or a filter press, the reprecipitated solid is squeezed (e.g. between two conveyor belts) to maximise the removal of liquids from the solid. Exiting the filter 670 is a cake 674 comprising the reprecipitated solid. The cake 674 typically comprises -40% solids by mass. The neutralised solution 672 exits the filter 670 as the filtrate. The filter 670 can comprise an optional wash stage. As above, the wash stage increases the recovery of the interstitial liquor from the cake.

[282] Since the neutralised solution 672 is now substantially free of the nitric acid-soluble components (e.g. Ti) and is substantially free of other impurities, the neutralised solution 672 is collected for reuse within the process. As described below, the neutralised solution 672 is typically combined with the solution 653 recovered from dewatering 651.

[283] The reprecipitated solid 674 is transported, e.g. by solids conveyor, to a dewatering centrifuge 651. The dewatering centrifuge 651 has the same configuration and operates in the same manner as described for dewatering centrifuge 308. From the solids outlet of the dewatering centrifuge 651, a substantially liquids-free reprecipitated solid 655 is recovered. From the centrate outlet of the dewatering centrifuge 651, liquor 653 entrained with the reprecipitated solid 674 that is removed therefrom by the dewatering centrifuge 651 is recovered.

[284] The composition of the liquor 653 is the same as the neutralised solution 672 recovered from the filter 670, i.e. liquor 653 is also a neutralised solution. The liquor 653 is therefore combined with the neutralised solution 672. The combined neutralised solution 902 is recovered for reuse within the process 10, as described in further detail below.

[285] The substantially liquids-free reprecipitated solid 655 is transported, e.g. by solids conveyor, to a dryer 676. In the dryer 676, the substantially liquids-free reprecipitated solid 655 is heated. The dryer 676 is typically an indirectly-heated rotary drum dryer, with the same configuration and operation as dryer 316. As above, the substantially liquids-free reprecipitated solid 655 from the dewatering centrifuge 651 enters the rotating drum of the rotary dryer 676 through a material inlet.

[286] As the substantially liquids-free reprecipitated solid 655 moves through the drum of the rotary dryer, energy transferred from the hot vapour to the solid via the internal walls and vanes of the dryer causes the solid to be dried. As the substantially liquids-free reprecipitated solid 655 is dried, water vapour is produced. In addition, as the titanium hydroxide is heated, it is decomposed back to titanium dioxide forming water, which also evaporates to form additional water vapour (see Reaction 4 below). Ti(OH) ₄ TiO ₂ + 2 H ₂ O (R4)

[287] The operating temperature of the rotary dryer 676 is selected to be above the decomposition temperature of titanium hydroxide, i.e. so that the titanium hydroxide decomposes to titanium dioxide. Titanium hydroxide can decompose at temperatures as low as about 100 °C. At this temperature, substantially all the water present in the reprecipitated solid 655, plus additional water generated by the decomposition of the titanium hydroxide, evaporates.

[288] It is advantageous to recover titanium as titanium dioxide because titanium dioxide is a useful raw material in many industries - e.g. conductors, pigments, textiles, paints etc.

[289] The dried solid comprising titanium dioxide and vapours exit the rotating drum at its exit. At the exit, the vapour stream 678 is separated from the dried titanium dioxide product 680.

[290] The vapour stream 678 separated from the dried solid at the discharge breech of the dryer 676 comprises water, entrained dried titanium dioxide product 680 and any other vaporised materials. When the concentration of dried solids in the vapour 678 is too high, the vapour is filtered (e.g. using a bag filter) to remove/recover entrained dried titanium dioxide product. The filtered vapour stream is then cooled to condense water present. Typically, the condensed water 678 is returned to the pond 102 when the pond 102 is active. When the pond 102 is inactive, the condensed water 678 can be used as process water elsewhere within the process.

[291] The recovered combined neutralised solution 902 is converted to saline water and nitric acid, as shown in Fig. 9. The neutralised solution 902 is transported, e.g. pumped, to a chlorination tank 904. In the chlorination tank 904, hydrochloric acid 906 is added. The hydrochloric acid 906 typically has a concentration of about 10% by mass hydrochloric acid.

[292] The solution 902 primarily comprises sodium nitrate. The solution 902 may also comprise sodium hydroxide in small quantities, e.g. because an excess of sodium hydroxide 666 is typically added to the precipitation tank 664. As the hydrochloric acid 906 is added to the chlorination tank 904, it neutralises any sodium hydroxide present, thereby forming sodium chloride and water.

[293] Because sodium chloride is soluble in water, the hydrochloric acid 906 and sodium nitrate in the solution 902 do not directly react. Instead, an aqueous solution 908 comprising sodium, chloride, hydronium and nitrate ions is formed. [294] The aqueous solution 908 is transported, e.g. pumped, to a heating vessel 910. In the heating vessel 910, the aqueous solution 908 is heated so as to boil off nitric acid present in the aqueous solution 908 (i.e. present as hydronium and nitrate ions). The heating vessel 910 operates at a temperature of about 85 °C, i.e. just above the boiling point of nitric acid. The heating vessel 910 is typically a jacketed vessel. Steam flows through the jacket. As the steam flows through the jacket, heat is transferred from the steam through the walls of the vessel to the aqueous solution 908 inside the vessel. The flow rate of steam through the jacket is adjusted so as to achieve the temperature of about 85 °C inside the vessel. At this temperature, primarily nitric acid boils off from the aqueous solution 908, as vapour 912. The water primarily remains liquid, i.e. because the temperature is below 100 °C, although a small amount may be stripped from the solution with the nitric acid vapour. Thus, the vapour 912 may comprise a small quantity of water vapour.

[295] The (primarily nitric acid) vapour 912 produced as a result of heating the aqueous solution 908 is captured, cooled and condensed 914. Cooling and condensation 914 may be performed by any suitable means known by the person skilled in the art, such as by using a heat exchanger with a suitable cooling medium. The vapour 912 is cooled to below 83 °C, so that it condenses, thereby forming a nitric acid condensate 916.

[296] The nitric acid condensate 916 is diluted 918 using recycled process water 920. The nitric acid is diluted to a concentration of approximately 68% by mass nitric acid (i.e. the azeotropic concentration). The diluted nitric acid 922 is transported, e.g. by pump, to a nitric acid storage tank, wherefrom it can be reused in the nitric acid digestion vessel 654.

[297] The solution remaining in the heating vessel 910 (i.e. after evaporation of the nitric acid) is a saline solution comprising primarily sodium chloride. The saline solution 924 is transported from the vessel 910, e.g. by pumping, for re-use either in the chlor-alkali plant or in the make-up of an eluant for the regeneration of one or more of the ion-exchange resins.

(iv) Mo separation and recovery - Fig. 6

[298] As illustrated in Fig. 6, the residual sludge 660 separated from the solution comprising nitric acid-soluble components is transported, e.g. by solids conveyor, to the Mo separation and recovery stage (iv) of the non-acid-soluble recovery area 600. In stage (iv), Mo (and any other dilute-alkali-soluble components present in the residual sludge 660) are recovered.

[299] The residual sludge 660 is added into (e.g. by solids conveyor or gravity) the dilute- alkali digester 682. In the dilute-alkali digester 682, a dilute solution of sodium hydroxide 684 is added to the residual sludge 660. The sodium hydroxide 684 has a concentration of about 1M (i.e. it is relatively dilute). The conditions of the dilute-alkali digester 682 are controlled so as to promote the dissolution of dilute-alkali-soluble components from the residual sludge. Typically, the dilute-alkali-soluble components will comprise Mo. However, depending on the composition of the impounded coal ash to the process, other dilute-alkali-soluble components may also be present.

[300] Mo present in the residual sludge 660 is in the form of molybdenum oxide (MoOs). As the ~ 1M sodium hydroxide 684 contacts the molybdenum oxide, the molybdenum oxide dissolves to form sodium molybdate (Na2MoO4), according to (Reaction 5):

MOO ₃ + NaOH Na ₂ MoO ₄ + H ₂ O (R5)

[301] Sodium molybdate is soluble in dilute-alkali solutions, so will remain in solution. Other oxides that may be present in the residual sludge 660 that are soluble in dilute-alkali solutions will likewise react with the sodium hydroxide. Components in the residual sludge that are not dilute-alkali-soluble will remain in the sludge as solids.

[302] The amount of about 1M sodium hydroxide 684 added to the dilute-alkali digester 682 is controlled so as to maximise the fraction of Mo recovered from the residual sludge. For example, a set volume/mass of 1M sodium hydroxide 684 can be added to the dilute- alkali digester 682 based on the mass of the residual sludge 660 and the concentration of Mo (and other dilute-alkali-soluble components) within the sludge, so as to ensure an excess of sodium hydroxide within the dilute-alkali digester 682. Alternatively, 1M sodium hydroxide 684 can be added to the dilute-alkali digester 682 so as to achieve a set pH within the dilute- alkali digester 682.

[303] The dilute-alkali digester 682 typically comprises some form of agitation, such as an agitator attached to a variable speed motor. Agitation of the contents of the dilute-alkali digester 682 ensures homogeneity within the dilute-alkali digester 682 and helps maximise the efficiency of the reaction between the molybdenum oxide in the residual sludge and the sodium hydroxide. The size of the dilute-alkali digester 682 is selected to produce sufficient residence time for the reaction between the molybdenum oxide and sodium hydroxide to achieve the maximum extent, thereby maximising recovery of Mo into solution.

[304] Exiting the dilute-alkali digester 682 is a slurry 686. The solid component of the slurry comprises a residual sludge comprising non-dilute-alkali-soluble components of the residual sludge 660. The liquid component of the slurry comprises a solution comprising the dilute- alkali soluble components, i.e. Mo. [305] The slurry 686 is passed, e.g. by a slurry pump (if a pumpable slurry) or by gravity, to a centrifuge bank 688. As above, the centrifuge bank 688 comprises one or more centrifuges which operates in parallel, with the slurry 686 being divided between the one or more centrifuges.

[306] As the slurry 686 passes into the centrifuge bank 688, the residual sludge 690 comprising non-dilute-alkali-soluble components is separated from the solution 692 comprising Mo. The residual sludge 690 has a solids fraction of about 50% by mass. As above, to further increase the recovery of the dilute-alkali-soluble components from the interstitial liquor of the residual sludge, the centrifuge bank 688 includes an optional wash stage.

[307] As above, in the wash stage, the residual sludge 690 is washed with fresh water or recycled liquor. As the residual sludge 690 is washed, the solution comprising the dilute- alkali-soluble components entrained within the solid residue is displaced. The displaced solution comprising the dilute-alkali-soluble components is recovered and added to the solution comprising Mo 692. The residual sludge 690 is transported, e.g. by solids conveyor, to the next stage (v) of the non-acid-soluble recovery area 600.

[308] The solution comprising Mo 692 is pumped, e.g. by a liquids pump, to an ionexchange stage 694. The ion-exchange stage 694 can operate as a batch, semi-batch or continuous process. It will be appreciated that the operation and number of columns will depend on the volume of solution 692 comprising Mo to be treated. For example, when the volume of the solution 692 comprising Mo is small, the ion-exchange stage 694 comprises a single packed bed column which operates in batch. On the other hand, when the volume of solution 692 comprising Mo is larger, the ion-exchange stage 694 may comprise two single packed bed columns which operate in parallel alternating between loading/regenerating, or stage 694 may comprise a loading and an elution column operating in series. Typically, the ion-exchange stage 694 operates as a batch process with a single column because the amount of Mo to be loaded is typically small.

[309] As above, the solution 692 comprising the Mo is pumped into the loaded column of the ion-exchange stage 694. The column comprises a resin which is specifically selected to preferentially react with Mo in the form of molybdate. As the solution comprising Mo is pumped through the loaded column, molybdate therein reacts with the resin and loads onto the resin. Exiting the column is the solution 696 which is substantially free of molybdate ions (and therefore Mo). [310] Because the solution 696 is substantially free of all impurities, it is recovered for reuse within the process 10. Typically, the solution 696 is used in the make-up of the 1M sodium hydroxide solution 684. For example, by adding an appropriate amount of concentrated sodium hydroxide to the solution 696 until it has a sodium hydroxide concentration of 1M.

[311] The resin (now loaded with Mo) is eluted. As above, the solution 692 stops being pumped into the column and a suitable eluant, such as sodium chloride or sodium hydroxide with a concentration of about 10% by mass, is instead pumped through the column, regenerating the resin. A concentrated (Mo) eluate 698 is pumped from the top of the column. The concentrated Mo eluate 698 (which comprises Mo in the form of MoOs) is added to the exotic metal alloy product 320 (see Fig. 2). This is because several valuable exotic metal alloys contain Mo in various quantities.

(v) Ge separation and recovery - Fig. 6

[312] The residual sludge 690 comprising non-dilute-alkali-soluble components separated from the solution comprising the dilute-alkali-soluble components is transported, e.g. by solids conveyor, to stage (v) of the non-acid-soluble recovery area 600. In stage (v), those components of the residual sludge 690 which are soluble in concentrated alkali solutions are separated and recovered.

[313] The residual sludge 690 is added into (e.g. by solids conveyor or gravity) the concentrated-alkali digester 601. In the concentrated-alkali digester 601, a concentrated solution 603 of sodium hydroxide is added to the residual sludge 690. The sodium hydroxide must be concentrated enough to cause dissolution of the concentrated-alkali-soluble components from the residual sludge. The sodium hydroxide solution 603 typically has a concentration of ~18M (i.e. it is relatively concentrated). The concentrated-alkali-soluble components typically comprise Ge. However, depending on the composition of the impounded coal ash to the process, other concentrated-alkali-soluble components can be present, and Ge may not be present, i.e. if the impounded coal ash did not comprise Ge. In this latter case, stage (v) can be isolated or omitted from the process.

[314] Ge in the residual sludge 690 is in the form of germanium oxide (GeO2). As the ~18M sodium hydroxide solution 603 contacts the germanium oxide, the germanium oxide dissolves to form sodium germanate (Na2GeO ₃ ), according to (Reaction 6):

GeO ₂ + 2 NaOH Na ₂ GeO ₃ + H ₂ O (R6) [315] Sodium germanate is soluble in concentrated-alkali solutions, so will remain in solution. Other oxides that may be present in the residual sludge 690 that are soluble in concentrated-alkali solutions will likewise react with the sodium hydroxide. Components in the residual sludge that are not concentrated-alkali-soluble will remain in the sludge.

[316] The amount of 18M sodium hydroxide 603 added to the concentrated-alkali digester 601 is controlled so as to maximise the fraction of Ge (and other concentrated-alkali-soluble components when present) recovered from the residual sludge. For example, a set volume/mass of 18M sodium hydroxide 603 can be added to the concentrated-alkali digester 601 based on the mass of the residual sludge 690 and the concentration of Ge (and other concentrated alkali-soluble components when present) within the sludge, so as to ensure an excess of sodium hydroxide within the concentrated-alkali digester 601. Alternatively, 18M sodium hydroxide solution 603 can be added to the concentrated-alkali digester 601 so as to achieve a set pH within the concentrated-alkali digester 601. The set pH is selected such that an excess of sodium hydroxide is present within the concentrated-alkali digester 601. Typically, the mass of 18M sodium hydroxide 603 added is at least equal to the mass of solids in the residual sludge 690.

[317] The concentrated-alkali digester 601 typically comprises some form of agitation, such as an agitator attached to a variable speed motor. Agitation of the contents of the concentrated-alkali digester 601 ensures homogeneity within the concentrated-alkali digester 601 and helps maximise the efficiency of the reaction between the germanium oxide in the residual sludge and the sodium hydroxide. The size of the concentrated-alkali digester 601 is selected so as to ensure sufficient residence time for the reaction between the germanium oxide and sodium hydroxide to achieve the maximum extent, thereby maximising recovery of Ge.

[318] Exiting the concentrated-alkali digester 601 is a slurry 605. The solid component of the slurry comprises a residual sludge comprising non-concentrated-alkali-soluble components of the residual sludge 690. The liquid component of the slurry comprises a solution comprising the concentrated-alkali-soluble components (i.e. Ge).

[319] The slurry 605 is passed, e.g. by a slurry pump or by gravity, to a centrifuge bank 607. As above, the centrifuge bank 607 comprises one or more centrifuges which operate in parallel, with the slurry 605 divided between the one or more centrifuges. As the slurry 605 passes into the centrifuge bank 607, the residual sludge 609 comprising non-concentrated- alkali-soluble components is separated from the solution 611 comprising the concentrated- alkali-soluble components. The residual sludge 609 comprises about 50% solids by mass. To increase the recovery of the concentrated-alkali-soluble components, the centrifuge bank 607 includes an optional wash stage.

[320] As above, in the wash stage, the residual sludge 609 is washed with fresh water or recycled liquor which comprises minimal impurities. As the residual sludge 609 is washed, the solution comprising the concentrated-alkali-soluble components entrained within the solid residue is displaced by the fresh liquor. The displaced solution comprising the concentrated-alkali-soluble components is recovered and added to the solution

611 comprising the concentrated-alkali-soluble components.

[321] The residual sludge 609 is transported, e.g. by solids conveyor, to the next stage (vi) of the non-acid-soluble recovery area 600.

[322] The solution 611 comprising the concentrated-alkali-soluble components is transported, e.g. by liquor pump, to a precipitation tank 613. In the precipitation tank 613, hydrochloric acid 615 is added to the solution 611 comprising the concentrated-alkali-soluble components. As the hydrochloric acid 615 is added to the precipitation tank 613, it reacts with the sodium hydroxide, thereby neutralising the solution. Further, as the hydrochloric acid 615 is added to the precipitation tank 613, the sodium germanate reacts with the hydrochloric acid, according to (Reaction 7):

Na ₂ GeO ₃ + 2 HCI + H ₂ O^ Ge(OH) ₄ + 2 NaCI (R7)

[323] Because germanium hydroxide is insoluble under neutral/acidic conditions, the germanium is caused to precipitate from the solution, as the insoluble hydroxide, i.e. because the solution in the precipitation tank is neutralised or becomes slightly acidic.

[324] Other concentrated-alkali-soluble components that are present in the solution 611 can similarly react with hydrochloric acid, forming insoluble precipitates.

[325] Hydrochloric acid 615 is added until the solution in the precipitation tank 613 is neutralised, i.e. it has a pH of about 7. The pH of the precipitation tank 613 is monitored, e.g. with the use of a pH probe, and hydrochloric acid 615 is added until the pH is about 7. This ensures that all the concentrated-alkali-soluble components in the solution 611 are precipitated.

[326] The precipitation tank 613 typically comprises some form of agitation, such as an agitator attached to a variable speed motor. Agitation of the contents of the precipitation tank 613 ensures homogeneity within the precipitation tank 613 and helps maximise the efficiency of the reaction between the sodium germanate in the solution 611 and the hydrochloric acid 615. The size of the precipitation tank 613 is selected to provide sufficient residence time for the reaction between the sodium germanate and the hydrochloric acid to achieve its maximum extent, thereby maximising recovery of Ge into the precipitated solid.

[327] Exiting the precipitation tank 613 is a slurry 617. The solid component of the slurry comprises the precipitated concentrated-alkali-soluble components (e.g. Ge) that reprecipitated in the precipitation tank 613. The liquid component of the slurry comprises a neutralised solution, which is substantially free of the concentrated-alkali-soluble components (e.g. Ge).

[328] The slurry 617 is pumped by a slurry pump to a filter 619. For example, the filter 619 can be a belt filter or a filter press. The type of filter 619 is selected depending on the volume of the slurry 617, the solids fraction of the slurry and the properties of the reprecipitated solid in the slurry. As the slurry 617 is filtered by the filter 619, the solid component of the slurry comprising the precipitated solid is separated from the liquid component of the slurry comprising the neutralised solution. When the filter 619 is, e.g. a belt filter or a filter press, the precipitated solid is squeezed (e.g. between two conveyor belts) to maximise the removal of liquids from the solid. Exiting the filter 617 is a filter cake 621 comprising precipitated solid and the neutralised solution 623. The filter cake 621 typically comprises -50% solids by weight.

[329] Since the neutralised solution 623 is now substantially free of the concentrated-alkali- soluble components (e.g. Ge) and is substantially free of other impurities, the neutralised solution 623 is collected for reuse within the process, as described below.

[330] The filter cake 621 is transported to a dewatering centrifuge 657. The dewatering centrifuge 657 operates in the same manner as dewatering centrifuge 308 (Fig. 2b). As above, as the filter cake 621 enters the dewatering centrifuge 657, the rotation of the screw conveyor and bowl of the dewatering centrifuge cause the solids and liquids within the filter cake 621 to separate. The solids are scraped towards the solids outlet of the dewatering centrifuge, forming a dewatered precipitated solid 661. The liquids move to the centrate outlet, forming a recovered liquor 659.

[331] The recovered liquor 659 has the same composition as the filtrate 623, i.e. it is a neutralised solution substantially free of Ge. The recovered liquor 659 is combined with the filtrate 623 and is reused within the process 10. Because the recovered liquors 623, 659 are saline (i.e. they primarily comprise sodium chloride), they can either be used to generate an eluant for the ion-exchange stages where a saline eluant is required or can be used as a feed to the chlor-alkali plant. [332] The dewatered precipitated solid 661 is transported, e.g. by solids conveyor, to a dryer 625. In the dryer 625, the dewatered solid 661 is heated. The dryer 625 is typically an indirectly-heated rotary drum dryer, of the type and function described with reference to Fig. 2c. As above, the dewatered solid 661 from the dewatering centrifuge 657 enters the rotating drum of the rotary dryer 625 through a material inlet.

[333] Combustion of gas in the burners of the rotary dryer provide thermal/heat energy in the form of hot vapours and heat the walls of the rotating drum. As above, energy from the hot vapours is transferred to the dewatered solid 661 through the internal walls and vanes of the rotary drum dryer. As the dewatered solid 661 moves through the drum of the rotary dryer, this energy transfer causes the solid to be dried. As the dewatered solid is dried, water vapour is produced. In addition, as the germanium hydroxide is heated, it is decomposed back to germanium oxide, forming water at the same time (see Reaction 8 below).

Ge(OH) ₄ GeO ₂ + 2 H ₂ O (R8)

[334] The operating temperature of the rotary dryer 625 is selected such that it operates above the decomposition temperature of germanium hydroxide, i.e. so that the germanium hydroxide decomposes to germanium oxide. The decomposition temperature of germanium hydroxide may be as low as 80 °C. The rate of decomposition is dependent on the reaction temperature, with higher reaction temperatures increasing the reaction rate. Typically, a temperature of about 100 °C is selected such that substantially all the water present in the dewatered solid, plus additional water generated by the decomposition of the germanium hydroxide is caused to evaporate, i.e. because the temperature in the rotary dryer will be set above 100 °C, to evaporate the water.

[335] It is advantageous to recover germanium as germanium oxide because germanium oxide is a useful raw material in many industries - e.g. semiconductors, catalysts, transistors etc.

[336] The dried solid comprising germanium oxide and water vapour exits the rotating drum at the breech exit. At the exit, the vapour stream 627 is separated from the dried germanium oxide product 629.

[337] The vapour stream 627 separated from the dried solid at the discharge breech comprises water, entrained dried germanium oxide product 629 and any other vaporised materials. When the concentration of entrained dried solid is too high, the vapour is filtered (e.g. using a bag filter) to remove/recover entrained dried germanium oxide product. The filtered vapour stream is then cooled to condense water present. Because the condensed water is essentially pure, it can be recycled back to the pond 102 when the pond 102 is active. Alternatively, when the pond 102 is inactive, the condensed water can be reused within the process 10 as process water.

(vi) Zr and Si separation and recovery- Fig. 6

[338] The residual sludge 609 separated from the solution comprising the concentrated- alkali-soluble components in the centrifuge bank 607, is transported, e.g. by solids conveyor, to the stage (vi) of the non-acid-soluble recovery area 600. The residual sludge 609 now primarily comprises Si and/or Zr - i.e. it is essentially comprised of inert materials that were not soluble in any of the dissolution media of stages (i) to (v).

[339] The residual sludge 609 passes to a dewatering stage 631. As the residual sludge 609 is passed to the dewatering stage 631, hydrochloric acid 665 is added thereto. The hydrochloric acid 665 neutralises any residual sodium hydroxide in the residual sludge 609. In particular, hydrochloric acid 665 is added to ensure the residual sludge 609 has a pH of between about 7 to about 8. This is because the LMEF product 637 and SLWA product 643 cannot be too alkaline. The addition of hydrochloric acid 665 is controlled by measuring the pH of the residual sludge 609 and incrementally adding hydrochloric acid 665 until the pH is within the range of 7-8.

[340] Typically, the dewatering stage 631 comprises a dewatering screw press conveyor, which operates in a similar manner to the dewatering centrifuge 308. Exiting the screw press conveyor at the solids outlet is the dewatered residual sludge 633. Exiting the screw press conveyor at the centrate outlet is a liquor 635 removed from the residual sludge 609 in the dewatering stage 631. Because the liquor 635 is substantially free of contaminants, it is recovered for reuse within the process 10. For example, the liquor 635 can be reused in the non-acid-soluble recovery area 600.

[341] The dewatered residual sludge 633 comprises a moisture content of around 20-40%. It will be appreciated that the exact moisture content is dependent on the method used to dewater the residual sludge 609 in the dewatering stage 631. The dewatered residual sludge 633 is suitable for sale and use as a lightweight manufactured earth fill (LMEF) 637 in road construction and for mine void fillings. Some or all of the dewatered residual sludge 633 can be sold for use as an LMEF 637.

[342] A remaining portion of the dewatered residual sludge 639 is transported, e.g. by solids conveyor or manually, to a dryer 641. Typically, the dryer 641 is an indirectly-heated rotary drum dryer, of the type and function described with reference to Fig. 2c. As above, the dewatered residual sludge 639 from the screw press 631 enters the rotating drum of the rotary dryer 641 through a material inlet. As the residual sludge 639 moves through the rotating drum and contacts the internal walls and vanes of the rotating drum, thermal energy is transferred to the residual sludge. This causes the residual sludge to be dried.

[343] The dried solid comprising Si and Zr and water vapour exit the rotating drum at the discharge breech. At the exit, the vapour stream 645 is separated from the dried Si/Zr product 643.

[344] The vapour stream 645 separated from the dried solid at the discharge breech comprises water, entrained dried Si/Zr product and any other vaporised materials. When the concentration of entrained dried Si/Zr product is too high, the vapour is filtered (e.g. using a bag filter) to remove/recover entrained dried Si/Zr product. The filtered vapour stream is then cooled to condense water present. Because the condensed water is essentially pure, a portion of the condensate is sent to the nitric acid dilution vessel 918, i.e. where it is needed to make up the concentrated nitric acid 922 added to the nitric acid digestion vessel 656. Excess condensate is pumped back to the pond 102, when the pond is active. When the pond is inactive, excess condensate is used elsewhere in the process 10 as process water.

[345] The dried Si/Zr product 643 exiting the dryer 641 is suitable for incorporation into a structural light weight aggregate, for example the structural light weight aggregate of AU 2021106969.

[346] Advantageously, the split of the dewatered residual sludge 633 to direct sale as an LMEF 637 and to further drying for use in a structural light weight aggregate 639 can be easily tailored. Initially, all the dewatered residual sludge 633 can be sold directly as LMEF, so that construction of the dryer 641 is avoided. Once the dryer 641 is constructed, the fraction of dewatered residual sludge 639 sent to the dryer 641 can be adjusted based on the material requirements of the structural light weight aggregate production area. For example, when more material is needed to produce the structural light weight aggregate, a larger fraction of the dewatered residual sludge is sent to the dryer.

Acid-Soluble Recovery Area 700 - Fig. 7

[347] The solution 502 comprising the acid-soluble components of the coal ash residue (see Fig. 4) - i.e. that is separated from the dewatered solid comprising the non-acid-soluble components and precipitated chlorides is pumped to the acid-soluble recovery area 700 (i.e. Fig. 7). Referring now to Fig. 7, in the acid-soluble recovery area 700, acid-soluble components 702 in the solution 502 are extracted and recovered therefrom. Typically, the acid-soluble components in the solution 502 comprise: Sb, Be, Cd, Zn, Sn. Of these, Cd and Zn are toxic elements and need to be converted to a safe/stable form.

[348] The acid-soluble recovery area 700 comprises one or more processing stages in which one or more of the acid-soluble components 702 are recovered as a product. It will be appreciated that the number of processing stages can be varied, depending on the number of acid-soluble components in the solution 502. This is ultimately dependent on the composition of the impounded coal ash being remediated.

[349] When the impounded coal ash comprises all the elements Sb, Be, Cd, Zn, Sn, there are five processing stages in which the elements are extracted and recovered from the solution 502. Typically, each of the processing stages comprise ion-exchange. However, it will be appreciated that other processes (chemical and/or physical) may also be used. Also, when not all the elements Sb, Be, Cd, Zn, Sn are present in the impounded coal ash, one or more of the processing stages can be bypassed, isolated or omitted.

[350] The solution 502 comprising the acid-soluble components 702 is pumped to a first ion-exchange stage 703. As above, the ion-exchange stage 703 may be operated as a semi- continuous or batch process using either a single or multiple column, depending on the volume of solution 502 to be treated and the concentration of the acid-soluble components in the solution 502. Typically, each of the ion-exchange stages 703, 708, 714, 720, 726 operate as batch processes with a single column.

[351] In the first-ion exchange stage 703, a resin is selected that preferentially reacts with Sb ions. A particularly suitable resin for Sb adsorption is a weak acid cation exchanger such as AmberLite IRC83 H Resin. As the solution 502 is pumped through the column of the first ion-exchange stage 703, Sb ions in the solution adsorb to the resin (i.e. by exchanging with other ions loaded onto the resin). A solution 704 substantially free of Sb is pumped from the top of the column of the first ion-exchange stage 703.

[352] The resin (now loaded with Sb) is eluted. As above, the solution 502 stops being pumped into the column and a suitable eluant, such as hydrochloric acid, is instead pumped through the column, regenerating the resin. The hydrochloric acid is produced by the chloralkali plant (see Fig. 10). A concentrated (Sb) eluate 706 passes from the top of the column during elution and is transported to a tank 732 for storage. Once the resin is regenerated, the pump pumping the eluant is stopped, and the column is again used for loading, i.e. by pumping solution 502 into the column.

[353] The solution 704 substantially free of Sb is pumped to the column of the second ionexchange stage 708. In the second ion-exchange stage 708, a resin is selected that preferentially reacts with Be ions. A particularly suitable resin for Be adsorption is a strong acid cation exchange resin, such as AmberLite IR120 or IRC86. As above, as the solution 704 is pumped through the column of the second ion-exchange stage 704, Be ions in the solution adsorb to the resin (i.e. by exchanging with other ions loaded onto the resin). A solution 710 substantially free of Be (and Sb) is pumped from the top of the column of the second ion-exchange stage 703.

[354] The resin (now loaded with Be) is eluted. As above, the solution 704 stops being pumped into the column and a suitable eluant, such as hydrochloric acid, is instead pumped through the column, regenerating the resin. A concentrated (Be) eluate 712 passes from the top of the column during elution and is transported to a tank 734 for storage. Once the resin is regenerated, the pump pumping the eluant is stopped, and the column is again used for loading, i.e. by pumping solution 704 into the column.

[355] The solution 710 substantially free of Be (and Sb) is pumped to the column of the third ion-exchange stage 714. In the third ion-exchange stage 714, a resin is selected that preferentially reacts with Cd ions. A particularly suitable resin for Cd adsorption is a strong base resin, such as AmberSep 21K XLT. As above, as the solution 710 is pumped through the column of stage 714, Cd ions in the solution 710 adsorb to the resin (i.e. by exchanging with other ions loaded onto the resin). A solution 716 substantially free of Cd (and Sb, Be) is pumped from the top of the column of the third ion-exchange stage 714.

[356] The resin (now loaded with Cd) is eluted. As above, the solution 710 stops being pumped into the column and a suitable eluant, such as sodium hydroxide, is then pumped through the column, regenerating the resin. A concentrated (Cd) eluate 718 passes from the top of the column during elution and is transported to a tank 736 for storage. Once the resin is regenerated, the pump pumping the eluant is stopped, and the column is again used for loading, i.e. by pumping solution 710 into the column.

[357] The solution 716 substantially free of Cd (and Sb, Be) is pumped to the column of the fourth ion-exchange stage 720. In the fourth ion-exchange stage 720, a resin is selected that preferentially reacts with Zn ions. A particularly suitable resin for Zn adsorption is a strong acid cation exchange resin such as AmberSep G26 H. As above, as the solution 716 is pumped through the column of stage 720, Zn ions in the solution 716 adsorb to the resin (i.e. by exchanging with other ions loaded onto the resin). A solution 722 substantially free of Zn (and Sb, Be, Cd) is pumped from the top of the column of the fourth ion-exchange stage 720. [358] The resin (now loaded with Zn) is eluted. As above, the solution 716 stops being pumped into the column and a suitable eluant, such as hydrochloric acid, is then pumped through the column, regenerating the resin. A concentrated (Zn) eluate 724 passes from the top of the column during elution and is transported to a tank 738 for storage. Once the resin is regenerated, the pump pumping the eluant is stopped, and the column is again used for loading, i.e. by pumping solution 716 into the column.

[359] The solution 722 substantially free of Zn (and Sb, Be, Cd) is pumped to the column of the fifth ion-exchange stage 726. In the fifth ion-exchange stage 726, a resin is selected that preferentially reacts with Sn ions. A particularly suitable resin for Sn adsorption is a strong base resin such as AmberSep 21K XLT. As above, as the solution 722 is pumped through the column of the fifth ion-exchange stage 726, Sn ions in the solution adsorb to the resin (i.e. by exchanging with other ions loaded onto the resin). A solution 728 substantially free of Sn (and Sb, Be, Cd, Zn) is pumped from the top of the column of the fifth ion-exchange stage 726.

[360] The resin (now loaded with Sn) is eluted. As above, the solution 722 stops being pumped into the column and a suitable eluant, such as sodium hydroxide, is then pumped through the column, regenerating the resin. A concentrated (Sn) eluate 730 passes from the top of the column during elution and is transported to a tank 740 for storage. Once the resin is regenerated, the pump pumping the eluant is stopped, and the column is again used for loading, i.e. by pumping solution 722 into the column.

[361] The (Sb, Be, Cd, Zn, Sn) concentrated eluates are periodically pumped from the respective storage tanks 732, 734, 736, 738, 740, for example when the storage tanks near capacity, for use as components in a cementitious product. The cementitious product can comprise, for example, the structural light weight aggregate described in AU 2021902692. When the concentrate eluates are used as a component of the structural light weight aggregate, the solutions from tanks 732, 734, 736, 738, 740 are pumped to the dryer prior to the particle separation stage of the process of AU 2021902692.

[362] However, depending on the concentrations of the (Sb, Be, Cd, Zn, Sn) concentrated eluates, one or more of the (Sb, Be, Cd, Zn, Sn) concentrated eluates may be suitable for use as a component of another type of cementitious product. For example, one or more of the concentrated eluates can be used as components of cement roof tiles, cement blocks and/or concrete panels.

[363] Advantageously, in the acid-soluble recovery area 700, the toxic acid-soluble components Zn and Cd are extracted from the aqueous solution 502 and used as components of a cementitious product. By using the toxic components in the cementitious product, they are stabilised - i.e. they are no longer in a state in which they can leach etc. into the surrounding environment.

[364] As set forth above, the ion-exchange stages (703, 708, 714, 720, 726) typically operate in a batch mode. It will be appreciated that the exact operating conditions of each of the ion-exchange stages (703, 708, 714, 720, 726), e.g. temperature, pressure, flow-rates, etc., will vary and will depend on the properties of the resin used.

[365] In this embodiment, the acid-soluble components are extracted from the solution 502 in the order: Sb, Be, Cd, Zn and Sn. However, it will be appreciated by those skilled in the art that, depending on the selection of the resin, the acid-soluble components may be extracted in a different order, for example by selecting different resins. As another example, the ionexchange stages comprising strong acid cation resins can be performed first and the ionexchange stages comprising strong base anion resins can be performed second.

Precipitation Area 800 - Fig. 8

[366] Referring now to Fig. 8, the solution 728 substantially free of Sn (and Sb, Be, Cd, Zn) is pumped from the fifth ion-exchange stage to the precipitation area 800. The solution 728 comprises Al, Ca and Mg in the form of chlorides. In the precipitation area 800, aluminium is recovered as alumina, and calcium and magnesium are recovered as dolomite and calcite.

[367] The solution 728 substantially free of Sn is pumped to the precipitation tank 803 of the precipitation area 800. In the precipitation tank 803, sodium carbonate 805 is added to the solution 728. The sodium carbonate 805 can either be in the form of sodium carbonate decahydrate crystals or can be a solution comprising sodium carbonate.

[368] As the sodium carbonate 805 is added to the precipitation tank 803, the chlorides of Al, Ca and Mg in the solution react with the sodium carbonate 805 to form alumina, calcium carbonate and magnesium carbonate respectively, according to the following reactions:

AI ₂ CI ₆ + 3 Na ₂ CO ₃ AI ₂ O ₃ + 6 NaCI + 3CO ₂ (R9)

CaCI ₂ + Na ₂ CO ₃ CaCO ₃ + 2 NaCI (R10)

MgCI ₂ + Na ₂ CO ₃ — > MgCO ₃ + 2 NaCI (R11)

[369] At the same time, the pH of the solution in the precipitation tank 803 increases (i.e. becomes more alkali). As the pH becomes more alkali, the alumina, calcium carbonate and magnesium carbonate start to precipitate from the solution, as they are insoluble under alkali conditions. Sodium carbonate 805 is added until substantially all the alumina, calcium carbonate and magnesium carbonate have precipitated from the solution. Typically, a pH of ~10 is required in the precipitation tank 803 to maximise precipitation of the alumina, calcium carbonate and magnesium carbonate.

[370] Carbon dioxide 806 produced by (Reaction 9) is captured, e.g. by scrubbing a vapour stream released from the precipitation tank 803, for re-use in the production of sodium carbonate and to reduce the carbon footprint of the process. For example, the recovered carbon dioxide can be bubbled through a solution comprising sodium hydroxide to regenerate a solution comprising sodium carbonate which can be added back to the precipitation tank 803.

[371] The precipitation tank 803 typically comprises some form of agitation to promote homogeneity within the precipitation tank 803. The size of the precipitation tank 803 is selected to produce sufficient residence time for the reaction between the sodium carbonate and the Al, Ca and Mg.

[372] The conditions of the precipitation tank 803 are controlled so as to minimise the formation of calcium aluminate precipitates, for example by keeping the temperature of the precipitation tank 803 at around ambient conditions. This is because calcium aluminate precipitates will form part of the residue 820, contaminating the alumina product. The calcium aluminate precipitate cannot be easily separated from the alumina product, reducing the purity of the alumina product. This, in turn, reduces the value of the alumina product.

[373] Exiting the precipitation tank 803 is a slurry 808 comprising a solid comprised substantially of alumina, calcium carbonate and magnesium carbonate. The slurry 808 is transported, e.g. by a slurry pump, to dissolution tank 810.

[374] In dissolution tank 810, cool hydrochloric acid 812 is added thereto. The hydrochloric acid 812 typically has a concentration of 35% by mass. The hydrochloric acid 812 reacts with free sodium carbonate in the dissolution tank 810, according to the following reaction, thereby neutralising the solution:

Na ₂ CO ₃ + 2 HCI 2 NaCI + H ₂ O + CO ₂ (R12)

[375] Additionally, the hydrochloric acid 812 reacts with the (precipitated) calcium carbonate and the (precipitated) magnesium carbonate, according to the following reactions:

CaCOs + 2 HCI CaCI ₂ + H ₂ O + CO ₂ (R13)

MgCO ₃ + 2 HCI MgCI ₂ + H ₂ O + CO ₂ (R14) [376] By adding cool hydrochloric acid 812, reactions between the precipitated alumina and the hydrochloric acid 812 are minimised, i.e. the alumina remains as a precipitated solid. The use of hydrochloric acid is further advantageous because the neutralisation reaction with sodium carbonate (Reaction 12) produces sodium chloride, which can be reused in the process, e.g. as feed to the chlor-alkali plant. Also, the dissolution reactions (Reaction 13) and (Reaction 14) respectively produce calcium chloride and magnesium chloride, allowing reprecipitation of Ca and Mg as carbonates according to (Reaction 10) and (Reaction 11), and as described below.

[377] Sufficient hydrochloric acid 812 is added to maximise the dissolution of calcium carbonate and magnesium carbonate whilst minimising the dissolution of alumina. The mass of hydrochloric acid 812 is controlled by targeting a pH within the dissolution tank 810. A pH probe (or other similar instrument) is used to measure the pH in the tank 810. Hydrochloric acid 812 is added (either manually or via automated control systems) until the measured pH in the tank 810 is the same as a predetermined pH. Typically, the predetermined pH is around 8.

[378] By precipitating alumina, calcium carbonate and magnesium carbonate, and then redissolving the calcium/magnesium carbonates, it is possible to produce an alumina product of higher purity compared with precipitating the alumina at first-instance. However, it will be understood by those skilled in the art, that the alumina product can instead be precipitated at first-instance (i.e. by adjusting the pH with sodium carbonate 805 such that only alumina precipitates).

[379] The carbon dioxide 814 emitted by (Reaction 12), (Reaction 13) and (Reaction 14) is captured, e.g. by scrubbing a vapour stream released from the dissolution tank 810, for reuse in the production of sodium carbonate and to reduce the carbon footprint of the process. As above, the captured carbon dioxide is bubbled through a solution comprising sodium hydroxide, thereby production a solution comprising sodium carbonate which can be reused in precipitation tank 803 and/or precipitation tank 840.

[380] The dissolution tank 810 typically comprises some form of agitation to promote homogeneity within the dissolution tank 810. The size of the dissolution tank 810 is selected to produce sufficient residence time for the reaction between the acid and the precipitated calcium carbonate and magnesium carbonate.

[381] Exiting the dissolution tank 810 is a slurry 816. The solid component of the slurry comprises a solid substantially comprising alumina, i.e. because the calcium carbonate and magnesium carbonate dissolved. The liquid component of the slurry comprises a solution comprising (the dissolved) calcium and magnesium.

[382] Although in this embodiment precipitation and dissolution are performed in separate tanks (802, 810 respectively), it will be appreciated that they can be performed in the same tank. When both processing steps are performed in the same tank, the solution 704 is pumped to the tank. A sufficient quantity of sodium carbonate is first added to the tank so as to cause precipitation of the alumina, calcium carbonate and magnesium carbonate. After a predetermined length of time has elapsed (i.e. such that the precipitation reactions are completed), hydrochloric acid is next added to the tank so as to redissolve the calcium/magnesium carbonates. Exiting the (combined precipitation/dissolution) tank is the slurry 816.

[383] The slurry 816 is transported, e.g. by a slurry pump, to a centrifuge bank 818. As above, the centrifuge bank 818 comprises one or more centrifuges, with the slurry 816 divided between the centrifuges based on the number of operational centrifuges and/or capacity of the centrifuges within the centrifuge bank 818.

[384] As the slurry 816 passes through the centrifuge bank 818, the solid 820 substantially comprising alumina is separated as a slurry from the solution 822 comprising calcium and magnesium. Optionally, the solid 820 is washed to maximise recovery of the solution comprising calcium and magnesium. Typically, the slurry comprising the substantially alumina solid 820 comprises 50% solids by mass (the remaining 50% being entrained liquids).

[385] The slurry 820 is transported, e.g. by conveyor or a suitable pump, to a dewatering screw press conveyor 824, which operates in a similar manner to the dewatering centrifuge 308. Exiting the screw press at the solids outlet is the dewatered solid 826. Typically, the dewatered solid 826 comprises -60-80% solids by mass. Exiting the screw press at the centrate outlet is the liquor 828 removed from the solid 820 in the screw press 824. The liquor 828 is collected for re-use within the process. For example, because the liquor 828 comprises sodium chloride, it can be recycled for reuse as an eluant in one or more of the ion-exchange stages. The liquor 828 is typically pumped to a storage tank wherefrom it is distributed when/as required.

[386] The dewatered solid 826 is transported, e.g. by conveyor, to a dryer 830 in which the moisture content of the dewatered solid 826 is further reduced. The dryer 830 is typically an indirectly-heated rotary dryer, of the type previously described. As above, the dewatered solid 826 enters the rotating drum of the rotary dryer 830 through a material inlet. As the dewatered solid 826 moves through the rotating drum and contacts the internal walls and vanes of the drum, energy is transferred to the dewatered solid 826, thereby causing it to be dried.

[387] The dried solid 802 comprising alumina and vapours exit the rotating drum at a the discharge breech. At the exit, the vapour stream 838 is separated from the dried alumina product 802, which is ready for sale.

[388] The vapour stream 838 separated from the dried solid comprising alumina comprises water, entrained dried alumina product and any other vaporised materials. When the concentration of entrained dried alumina is too high, the vapour is filtered (e.g. using a bag filter) to remove/recover entrained dried alumina product. The filtered vapour stream is then cooled to condense water present. The condensate is pumped to the pond 102 when the pond is active. When the pond is inactive, the condensate is used elsewhere in the process

10 as process water.

[389] The solution 822 comprising calcium and magnesium separated from the slurry 820 by the centrifuge bank 818 is pumped, e.g. by a liquor pump, to a second precipitation tank 840. Sodium carbonate 842 is also added to the second precipitation tank 840. The sodium carbonate 842 acts to increase the pH (i.e. making it more alkali) in the precipitation tank 840.

[390] As the sodium carbonate 842 is added to the precipitation tank 840, the chlorides of Ca and Mg in the solution in the precipitation tank 840 react with sodium carbonate 842 to form calcium carbonate and magnesium carbonate, i.e. as per (Reaction 10) and (Reaction 11).

[391] The sodium carbonate also neutralises any hydrochloric acid remaining in solution, producing carbon dioxide 862. As above, the carbon dioxide 862 is scrubbed, e.g. with a solution comprising sodium hydroxide, to generate a sodium carbonate solution which can be reused in precipitation tank 803 and/or precipitation tank 840.

[392] Sodium carbonate 842 is added until a pH of the solution in the precipitation tank 840 is reached at which the calcium carbonate and magnesium carbonate precipitate from the solution. Typically, the calcium carbonate and magnesium carbonate precipitate as dolomite and calcite.

[393] Exiting the precipitation tank 840 is a slurry 844. The solid component of the slurry is a solid comprising the precipitated dolomite and calcite. The liquid component of the slurry is a solution substantially free of Al, Ca and Mg. [394] The slurry 844 is pumped, e.g. by a slurry pump, to centrifuge bank 846. As above, the centrifuge bank 846 comprises one or more centrifuges, with the slurry 844 divided between the centrifuges based on the number of operational centrifuges and/or capacity of the centrifuges within the centrifuge bank 846.

[395] As the slurry 844 passes through the centrifuge bank 846, the solid 848 comprising dolomite and calcite is separated from the solution 850 substantially free of Al, Ca and Mg. The solid 848 comprising dolomite and calcite is separated as a slurry comprising 50% solids by mass (the remaining 50% being liquids).

[396] The solution 850 substantially free of Al, Ca and Mg is collected for reuse within the process. For example, because the solution 850 comprises sodium chloride, it can be added to the same storage tank as the solution 828 for use as an eluant in the regeneration of ionexchange resin(s) or can be used as a feed into the chlor-alkali plant.

[397] The slurry 848 is transported, e.g. by conveyor, to a screw press 864. The screw press 864 operates as previously described. Exiting the screw press at the solids outlet is the dewatered solid 852. The dewatered solid 852 typically comprises -60-80% solids by mass. Exiting the screw press at the centrate outlet is the liquor 854 removed from the solid 852 in the screw press 864. The liquor 854 is collected for reused within the process. For example, because the liquor 854 comprises sodium chloride, it can be recycled for reuse as an eluant in one or more of the ion-exchange stages or it can be used in the chlor-alkali plant. The liquor 854 is typically pumped to the same storage tank as liquor 850 from the centrifuge bank 846, i.e. because both liquors comprise sodium chloride.

[398] The solid 852 is transported, e.g. by conveyor, to a dryer 856 in which the moisture content of the solid 852 is further reduced. The dryer 856 is typically an indirectly-heated rotary dryer, of the type previously described. As above, the dewatered solid 852 enters the rotating drum of the rotary dryer 856 through a material inlet. As the dewatered solid 852 moves through the rotating drum and contacts the internal walls and vanes of the drum, energy is transferred thereto, causing the dewatered solid to be dried.

[399] The dried solid comprising dolomite and calcite and vapours exits the rotating drum at the discharge breech. At the exit, the vapour stream 860 is separated from the dried dolomite and calcite product 804.

[400] The vapour stream 860 separated from the dried solid comprising dolomite and calcite comprises water, entrained dried dolomite and calcite product and any other vaporised materials. When the concentration of entrained dried dolomite and calcite product is too high, the vapour is filtered (e.g. using a bag filter) to remove/recover entrained dried dolomite and calcite product. The filtered vapour stream is then cooled to condense water present. The condensate is pumped back to the pond 102, when the pond is active. When the pond 102 is inactive, the condensate is used as process water elsewhere within the process 10.

[401] Depending on the composition of the impounded coal ash, the dried dolomite and calcite product 804 comprises varying quantities of calcium and magnesium. The dried dolomite and calcite product 804 can be further treated to separate a high magnesium component (i.e. primarily dolomite) from a lower magnesium component (i.e. primarily calcite). The high-Mg and low-Mg carbonates can be sold as separate products.

Alternatively, the dried dolomite and calcite product 804 can be directly sold as a product. As another alternative, the solid comprising dolomite and calcite 804 can also be directly used as a component of a structural light weight aggregate or other cementitious product.

[402] The end-use of the dried dolomite and calcite product 804 depends on several factors including: the concentration(s) of Mg and Ca; the existence of a market for the particular carbonate concentration produced; the need for raw materials in the structural light weight aggregate production area, etc.

Chlor-alkali plant 950 - Fig. 10

[403] Because the process 10 uses primarily hydrochloric acid as a source of acid, sodium hydroxide as a base, and requires sodium chloride (i.e. for regenerating ion-exchange resins), the process 10 can conveniently include a chlor-alkali plant 950. In the chlor-alkali plant 950, hydrochloric acid, sodium hydroxide and a sodium chloride solution are each produced, which are all used as reagents in various parts of the process 10.

[404] Referring now to Fig. 10, a chlor-alkali plant 950 is shown in which a sodium chloride solution 952 enters a first compartment 954 of a membrane cell 956. The sodium chloride solution 952 is typically a saturated brine having a concentration of approximately 26% by mass NaCI. The sodium chloride solution comprises the saline solutions recovered from other parts of the process 10, as hitherto described.

[405] Depending on the concentration of the sodium chloride solution(s) recovered from the process 10, the chlor-alkali plant 950 can include an optional concentration stage (not shown in Fig. 10). In the optional concentration stage, the sodium chloride solution(s) is concentrated so as to produce a solution with a concentration of -26% NaCI. For example, the concentration can be performed by reverse osmosis (RO) or evaporation. [406] The first compartment 954 of the cell 956 is separated from a second compartment 957 of the cell 956 by a membrane 958. The first compartment 954 has an electrode 960, which takes on a positive charge (i.e. acts as the anode) when an electrical current is caused to flow through the cell 956. The second compartment 957 has an electrode 962, which takes on a negative charge (i.e. acts as the cathode) when an electrical current is caused to flow through the cell 956.

[407] The membrane 958 is configured such that it is permeable to sodium cations (Na ⁺ ), but impermeable to other ions, including chloride (Cl’) and hydroxide (OH-).

[408] Water 961 is added to the second compartment 957 of the cell 956. At least a portion of the water 961 is typically fresh water, whilst another portion of the water can be recycled process water from within the process 10, e.g. process condensates from any one of the dryers.

[409] An electrical current is caused to flow through the cell 956 when the two electrodes 960, 962 are connected to an electrical source (not shown) and the electrical source is on. As the sodium chloride solution 952 enters the cell which is under the influence of an electrical current, chloride ions are oxidised to chlorine gas at the anode 960. Because the chlorine gas is less dense than the surrounding solution, it moves to the top of the cell 956 and is collected from a chlorine gas outlet 964.

[410] At the cathode 962, water is reduced by the electrons provided by the flow of electricity through the cell, forming hydrogen gas and hydroxide ions. Because the hydrogen gas is less dense than the surrounding solution, it moves to the top of the cell 956 and is collected from a hydrogen gas outlet 966.

[411] The hydrogen gas 966 and chlorine gas 964 are collected and used to produce hydrochloric acid for use in the process 10. For example, the hydrogen gas 966 and chlorine gas 964 can be directed into the same vessel wherein the hydrogen and chlorine react to form hydrogen chloride. The resulting hydrogen chloride is absorbed in water (e.g. recycled process water) to produce a hydrochloric acid solution. The hydrochloric acid solution can then be diluted as needed.

[412] Because the membrane 958 is permeable to sodium cations, the sodium cations tend to migrate from the first compartment 954 to the second compartment 957, i.e. because the sodium cations are positively charged they tend to move toward the cathode. In the second compartment 957, the sodium cations react with the hydroxide anions to form sodium hydroxide. A solution comprising the sodium hydroxide 968 is collected from an outlet of the second compartment 957. [413] The solution 968 comprising sodium hydroxide is used in the process 10, as hitherto described. Depending on the concentration of the solution 968 comprising sodium hydroxide required, further treatment stages, e.g. evaporation, concentration, etc., may be required, for example to generate the 18M sodium hydroxide required in digester 601 in the non-acid- soluble recovery stage 600.

[414] From an outlet of the first compartment 954, a sodium chloride solution 970 is collected. The sodium chloride solution 970 typically has a concentration of around 24% by mass sodium chloride, i.e. it is less concentrated than the solution 952 which enters the cell 956.

[415] The sodium chloride solution 970 can be used in the process 10, for example as an eluant in the regeneration of ion-exchange resins, as hitherto described. Alternatively, the sodium chloride solution 970 can also be concentrated, e.g. by reverse-osmosis, to a saturated solution, for reuse as (at least a portion of) the sodium chloride solution 952.

[416] Examples a. Non-limiting Examples of the process 10 and various stages thereof will now be provided.

[417] Example 1

[418] In this Example, a process for remediating coal ash to recover alumina and other valuable by-products is described. The coal ash for remediation was removed from the active coal ash pond as a slurry. The coal ash removed from the coal ash pond was primarily comprised of 33 elements.

[419] The coal ash was analysed to determine the form and quantity of the 33 elements comprising the coal ash. The results are given in Table 1 (in solids equivalent ppm).

[420] Table 1. Form and quantity (in solids equivalent ppm) of the 33 elements comprising the majority of the # coal ash feed to pond 102 and ^A impounded coal ash to process 10.

*carbon is found in the organic/unburnt particulate matter mixed with the coal ash. Mean concentration is of Coal Ash in NSW Power Stations as reported in the ADAA Handbook, 2 ^nd Edition 2013

[421] It was noted that, because the coal ash comprised all 33 elements accounted for in process 10, all areas/stages of the process 10 were required to fully remediate the coal ash.

[422] It was further noted that the coal ash slurry was produced in essentially the same manner as outlined in AU 2021902692 because the coal ash pond was an active pond.

[423] The inventor also noted that the materials of construction for the equipment of the process were corrosion resistant to acids and alkalis, due to the use of highly corrosive substances such as hydrochloric acid. Areas with high friction, such as all moving parts and pipes, were also made to be abrasion resistant by installing rubber or ceramic linings/coatings of pumps and pipes where required.

Impounded Pond Ash Removal

[424] The impounded coal ash was dredged from the active pond 102 (see Fig. 2) by means of a hydraulic dredge installed on a floating hydraulic dredge barge. Because the pond 102 was active, the hydraulic dredge removed a coal ash slurry 104. The slurry 104 was dredged at a rate of 3600 solid tpd (tonnes per day) and 7200 liquid tpd. Further water addition 106 to the slurry was not required, because the slurry 104 was already a pumpable slurry. However, the inventor noted that additional water 106 could be added to the slurry 104 if the slurry 104 was not a pumpable slurry.

[425] The slurry 104 was pumped to a trash screen 110 which was in the form of a conveyor with perforations. As the slurry 104 passed over the trash screen 110, non-ash waste 112 with particle sizes greater than the perforations on the trash screen 110 did not pass through the trash screen 110. The non-ash waste 112 was observed to comprise vegetative waste and tree roots. It was collected and sent to a compositing facility. The inventor noted that the mass of non-ash waste 112 would vary depending on e.g. whether the pond 102 was active or not. [426] The finer particles of the slurry 114 were collected as a clean ash slurry 114 and were pumped by a slurry pump to the material grading area 200.

Material Grading Area

[427] Fig. 2 shows the major processes employed to grade the clean ash slurry 114. In the setup of Fig. 2, clean ash slurry 114 was fed at a rate of approximately 10800 tpd to a flotation stage 202. In the flotation stage 202, the organic matter (i.e. carbon) and fibrous asbestos were removed from the clean ash slurry 114. The inventor noted that the solids dredged from the pond comprised around 1% organic matter (i.e. 36 tpd when 3600 solid tpd of impounded coal ash was dredged). The organic matter and fibrous asbestos were recovered as an overflow 204 from the flotation cell 214 of the flotation stage 202. The separation of the organic matter and fibrous asbestos was enhanced by sparging air through the flotation cells and by the addition of frothing agents.

[428] The overflow 204 comprising the organic matter and fibrous asbestos was passed from the flotation stage 202 to a nearby power station for combustion in the boiler burners as biomass. The overflow 204 comprised 36 solid tpd and was observed to be composed of carbon - i.e. all the carbon present in the clean ash slurry 114 was recovered to the overflow 204. The inventor noted that liquor from the overflow 204 was typically separated from the solid component and recycled back to the flotation cell.

[429] The carbon-free ash slurry 206 was discharged as an underflow from the bottom of the flotation cell 214 at a rate of 3564 solid tpd and 7200 liquid tpd. The carbon-free ash slurry 206 was pumped by a slurry pump to a coarse filtration stage 208, in which the coarser material was removed therefrom using an inclined static screen filter. The coarser material 210 was collected at the bottom of the inclined static screen filter at a rate of 108 solid tpd. The collected coarser material 210 was used to form a granular fill product (i.e. as described in AU 2021902692).

[430] The finer particles and liquids passed through the perforations of the static screen filter and collected as a treated ash slurry 212. A total of 10656 tpd of treated ash slurry 212 was produced, comprising 3456 solid tpd and 7200 liquid tpd. The treated ash slurry 212 was pumped by slurry pump to the drum magnetic separator 302 of the magnetic separation stage 300.

Magnetic Separation Stage

[431] As the treated ash slurry 212 was passed into the feed channel 302-1 of the drum magnetic separator 302 (see Fig. 2a), the magnetisable components of the treated ash slurry were attracted to the rotating shell 302-2. The non-magnetisable components 304 were not attracted to the rotating shell and were collected at a first outlet. The non- magnetisable components 304 were collected at a rate of 10433 tpd (comprising 3363 solid tpd and 7070 liquid tpd), i.e. substantially all the liquid exited the drum magnetic separator 302 at the first outlet.

[432] The magnetisable components 306, however, continued rotating with the shell 302-2 past the first outlet and were collected at a second outlet at a rate of 93 solid tpd and 130 liquid tpd. It was observed that the magnetisable components 306 had a sludge-like consistency, i.e. because the sludge still had a liquids content of -42% by mass. An analysis of the magnetisable components 306 showed it primarily comprised: FesC , O2O3, CoO, CuO, MnC>2, NiO, WO2 and V2O5, i.e. all the magnetic components in the impounded coal ash.

[433] The sludge of magnetisable material 306 was passed to a dewatering centrifuge 308. In the dewatering centrifuge 308, liquids were removed from the sludge by exerting a centrifugal force on the sludge. The removed liquids 312 were collected at a rate of 74 liquid tpd and pumped back to the pond 102. It was observed that the removed liquids 312 comprised very few solids.

[434] At the same time as liquids were removed from the sludge, a solid with a reduced moisture content was collected from a solids outlet 308-10 of the dewatering centrifuge 308. The solid 310 exiting the dewatering centrifuge 308 was observed to have a moisture content of around -38% by mass. To remove remaining liquor, the solid 310 was transported to an indirectly heated rotary drum dryer 316.

[435] In the rotary drum dryer 316, the solid 310 was heated as it contacted the internal walls and vanes of the dryer 316 which were externally heated from the combustion of natural gas. As the solid 310 was heated, the remaining liquids were evaporated and a dried metal alloy product 320 was produced at a rate of 93 solid tpd. The product 320 was observed to comprise negligible liquids, because the temperature and residence time of the rotary drum dryer was controlled so as to ensure substantially all the liquid was evaporated. The dried metal alloy product 320 was sold as a high-performance exotic metal alloy mix to a specialist steel foundry.

[436] The vapour was collected and cooled to condense water present. Approximately 56 tpd of condensed water was collected and pumped back to the pond 102.

Water-Soluble Component Recovery Area [437] The non-magnetisable components 304 collected from the drum magnetic separator 302 at the first outlet were pumped by slurry pump to the water-soluble component recovery area. Because most of the liquids were also collected at the first outlet, the inventor observed that the non-magnetisable components 304 were recovered as a slurry.

[438] Fig. 3 shows the major processes employed to recover the water-soluble components from the slurry 304. The inventor noted that, by this stage in the process 10, the water-soluble components of the impounded coal ash had dissolved from the coal ash. Therefore, the liquids in the slurry 304 comprised the water-soluble components. The remaining solid in the slurry 304 comprised the non-water-soluble components.

[439] In the set-up of Fig. 3, the slurry 304 was fed to a dewatering centrifuge 401 , which enabled a solution 406 comprising the water-soluble components to be recovered from the centrate outlet of the dewatering centrifuge 401. The solution 406 comprising the water- soluble components was recovered at a rate of 3774 liquid tpd. The inventor observed that the solution 406 comprised a negligible concentration of solids.

[440] An ash sludge 404 was also recovered from the dewatering centrifuge 401 at a mass flow rate of 6592 tpd (comprising 3296 liquid tpd and 3296 solid tpd). The ash sludge 404 was observed to comprise the non-water-soluble components and was transported to the acid digestion vessel 503.

[441] The inventor noted that the solution 406 comprising the water-soluble components comprised: As, B, Se, Ba, Li, K, Na.

[442] The solution 406 comprising the water-soluble components was passed through seven ion-exchange stages 408, 412, 420, 430, 440, 448, 456. Each ion-exchange stage 408, 412, 420, 430, 440, 448, 456 comprised a single column and operated as a batch process. The resins selected for each ion-exchange stage were selected so that the water- soluble components were recovered in the order: As, B, Se, Ba, Li, K, Na.

[443] The recovered water-soluble components As, B, Se, Ba, Li, K were recovered in the form of concentrated eluates (i.e. from the regeneration of the ion-exchange resins) and stored in respective storage tanks 415, 424, 434, 444, 452, 464. The concentrated eluate 460 comprising Na, however, was pumped directly to the chlor-alkali plant (see Fig. 10).

[444] The inventor noted that all the water-soluble components of the coal ash dredged from the pond 102 were able to be recovered in the water-soluble recovery area 300. Therefore, 0.04 solid tpd of arsenic oxide, 0.36 solid tpd of boron oxide, 0.01 solid tpd of selenium dioxide, 0.91 solid tpd of barium oxide, 0.5 solid tpd of lithium hydroxide, 52 solid tpd of potassium hydroxide and 13.9 solid tpd of sodium hydroxide were recovered from each of the seven ion-exchange stages 408, 412, 420, 430, 440, 448, 456 respectively.

[445] The inventor noted that As, B and Se were toxic, but were suitable to be incorporated as components into cementitious products. In particular, they were suitable to be incorporated into a structural light weight aggregate, as described in AU 2021902692. The concentrated eluates 414, 422, 432 were transported from the storage tanks 415, 424, 434 to the dryer prior to the particle separation stage of the process of AU 2021902692. This was because the concentrated eluates 414, 422, 432 were observed to still comprise liquids so had to be further dried to remove all moisture therefrom.

[446] On the other hand, the inventor noted that Ba, Li and K could be sold as valuable byproducts of the process 10. The concentrated eluates of Ba, Li, K could be sold directly from the storage tanks 444, 452, 464 respectively. However, because the concentrated eluates were in the form of solutions comprising Ba, Li, K, the eluates were removed from the storage tanks 444, 452, 464 and were dried in individual dryers to remove the liquids and produce dried Ba, Li, K products respectively.

Acid Leaching Area

[447] The ash sludge 404 was transported and added to the acid digestion vessel 503 (Fig. 4). The digestion vessel 503 was a continuously stirred tank reactor with a jacket through which steam was passed to provide heating thereto. The digestion vessel 503 was constructed of Hastelloy steel, because of the highly aggressive chemical environment and elevated temperatures. The inventor noted that, although in this particular example the digestion operated in a continuous mode, the digestion could also be performed as a batch process.

[448] Hydrochloric acid 504 with a concentration of 10% by mass was added to the digestion vessel 503 at a rate of 3296 tpd, i.e. acid was added at a 1:1 ratio by mass to the solids entering the vessel. The temperature of the digestion vessel 503 was measured and maintained at 90 °C by adjusting the flow of stream through the jacket. The elevated temperature promoted the dissolution of those components of the ash sludge 504 that were soluble in acid. At the same time, the use of hydrochloric acid promoted the conversion of metal oxides to metal chlorides. It was observed that some of the metal chlorides were insoluble and therefore formed a solid precipitate. Meanwhile, the metal oxides/chlorides that were soluble in an acidic environment formed a solution comprising the acid-soluble components. [449] A slurry 506 was removed from the digester which comprised a solution comprising the acid-soluble components of the ash sludge and a solid comprising the non-acid-soluble components and the precipitated chlorides. The slurry 506 was transported to a dewatering centrifuge 508, in which the solution 502 comprising the acid-soluble components was recovered from the centrate outlet at a rate of 4219 tpd (including 924 tpd of dissolved solids). The solution 502 was pumped to the acid-soluble recovery area 700.

[450] At the same time, an ash sludge 505 comprising the non-acid-soluble components and precipitated chlorides was recovered from the first outlet at a rate of 4745 tpd, consisting of 2372 solid tpd and 2372 liquid tpd. The ash sludge 505 was transported to the non-acid- soluble recovery area 600.

Non-Acid-Soluble Component Recovery Area

[451] The inventor noted that the ash sludge 505 comprising the non-acid-soluble components and precipitated chlorides was comprised of: Ge, Pb, Hg, Au, Ag, Ti, Mo, Si, Zr.

[452] The ash sludge 505 comprising the non-acid-soluble components and precipitated chlorides was added to the dissolution vessel 602. The dissolution vessel 602 was a continuously stirred tank reactor with a steam jacket. Sodium hydroxide 606 at a concentration of 10% by mass was added to the dissolution vessel 602 so as to neutralise acid present in the ash sludge 505. Hot water 604 at a temperature of 90 °C was also added to the vessel 602, which was maintained at a temperature of 90 °C by means of the steam jacket. The temperature within the vessel 602 was measured by a temperature probe and was used to adjust the mass flow of steam through the jacket, e.g. the mass flow was increased when the measured temperature was below 90 °C. The inventor noted that this was able to be achieved through the use of process automation.

[453] It was observed that as the sodium hydroxide 606 and hot water 604 were added to the vessel 602, HgCh and PbCl2 present in the sludge 505 was dissolved therefrom, because they were soluble in hot, neutral solutions. The total mass of sodium hydroxide 606 and hot water 604 added to the vessel 602 was 2372 tpd.

[454] The slurry 608 exiting the dissolution vessel 503 was observed to comprise a solution comprising Hg and Pb as chlorides and an ash sludge comprising the non-acid-soluble components and the remaining chlorides.

[455] The slurry 608 was passed to a centrifuge bank 610 which enabled the solution 614 comprising Hg and Pb as chlorides to be separated from the remaining ash sludge 612. The solution 614 was recovered at a rate of 2372 tpd. It was observed that the solution 614 comprised only a negligible concentration of solids. At the same time, the ash sludge was recovered as a slurry 612 comprising 50% by mass solids, with a total mass flow of 4744 tpd, i.e. 2372 solid tpd and 2372 liquid tpd.

[456] To recover the Hg and Pb from the solution 614, the solution 614 was passed through one ion-exchange stage 616. The ion-exchange stage 616 comprised a single column loaded with resin which operated in a batch mode. Resin was selected such that Pb and Hg were loaded onto the resin. The Pb and Hg were recovered from the ion-exchange stage as a concentrated eluate 624 at a rate of 1.0 tpd, which comprised 0.105 tpd of Pb and 0.0002 tpd of Hg.

[457] The inventor noted that both Pb and Hg were toxic, but the mass produced was small. Therefore, they could be stabilised by incorporation into a cementitious product, such as a structural light weight aggregate.

[458] A solution 649 substantially free of Pb and Hg was recovered at a rate of 2371 tpd. The inventor noted that the solution 649 could be reused in the chlor-alkali plant or for regenerating loaded resin (i.e. as an eluant).

[459] The ash sludge 612 recovered from the centrifuge bank 610 was transported to the digestion vessel 626. The vessel 626 was a continuously stirred tank reactor. A solution 628 comprising sodium thiosulphate at a concentration of 10% by mass was added at a rate of 2372 tpd to the ash sludge 612 in the vessel 626. This caused Au and Ag to be leached from the sludge, forming a solution comprising Au and Ag within the vessel 626.

[460] The slurry 634 from the digestion vessel 626 was passed to a centrifuge bank 636, enabling the recovery of the solution comprising Au and Ag 640. The solution 640 was recovered at a rate of 2372 tpd. At the same time, the remaining ash sludge (now substantially free of Au and Ag) was recovered as a slurry 638 comprising 50% by mass solids, with a total mass flow of 4744 tpd (i.e. 2372 solid tpd and 2372 liquid tpd).

[461] The solution 640 was pumped to a gold electrolysis chamber 642, which enabled the recovery of gold from the solution. In the chamber 642, the solution 640 was subjected to a current under which gold was caused to deposit onto carbon cathodes. Because the impounded coal ash used for this process comprised only trace amounts of gold, it was observed that the cathodes only needed to be replaced/scraped annually.

[462] The gold electrolysis chamber 642 was operated as a batch process, i.e. the solution 640 was held in the chamber 642 for a predetermined period of time. After the predetermined period of time, the remaining solution (now substantially free of Au) 646 was pumped to a silver electrolysis chamber 648.

[463] In the chamber 648, the solution 646 was subjected to a current under which silver was caused to deposit onto carbon cathodes. Because the impounded coal ash used for this process comprised only trace amounts of silver, it was observed that the cathodes only needed to be replaced/scraped annually.

[464] The silver electrolysis chamber was operated as a batch process, i.e. the solution 646 was held in the chamber 648 for a predetermined period of time. After the predetermined period of time, the remaining solution (now substantially free of Ag, as well as Au) 652 was recovered at an (average) rate of 2372 tpd.

[465] The solution 652 was used to generate the solution comprising sodium thiosulphate 628 added to the digestion vessel 626, i.e. by adding sodium thiosulphate to the solution 652 until it had a sodium thiosulphate concentration of 10% by mass.

[466] The inventor noted that because the solution 652 was continuously being recycled within the process 10, the concentration of any impurities in the solution 652 would build up over time. When the concentration of the impurities was above an upper limit, some of the solution 652 was instead directed back to the pond 102, i.e. it was not recycled. When this occurred, additional fresh water was required to produce the solution comprising sodium thiosulphate.

[467] The remaining ash sludge (now substantially free of Au and Ag) recovered as a slurry 638 from the centrifuge bank 636 was passed to a nitric acid digester 654. The nitric acid digester was a continuously stirred tank reactor.

[468] In the nitric acid digester 654, concentrated nitric acid 656 was added at a rate of 2372 tpd. The nitric acid 656 had a concentration of 68% by mass. The addition of the nitric acid 656 caused Ti to dissolve from the ash sludge to form titanium nitrate. Because titanium nitrate was soluble in concentrated nitric acid, it remained in solution.

[469] The resultant slurry 656 from the nitric acid digester 654 was pumped to a centrifuge bank 658, enabling a solution 662 comprising Ti to be separated. The solution 662 had a mass flow of 2400 tpd. It was observed that the solution 662 did not contain any solids. At the same time, the remaining ash sludge (now substantially free of Ti) was separated as a slurry 660 comprising 50% by mass solids, at a total mass flow of 4690 tpd (i.e. 2345 solid tpd and 2345 liquid tpd). [470] The solution 662 comprising Ti was pumped to a precipitation tank 664 where sodium hydroxide 666 at a concentration of 45% by mass was added at a rate of 2961 tpd. The addition of the sodium hydroxide 666 neutralised the solution 662 and caused Ti to precipitate as titanium hydroxide.

[471] To enable recovery of the precipitated titanium hydroxide, the resultant slurry 668 from the precipitation tank 664 was passed through a series of solid/liquid separation stages including a filter 670 and a dewatering centrifuge 651. Spray water was added to the filter 670 at a rate of 5 tpd to help rinse the precipitated titanium hydroxide of remaining interstitial liquor.

[472] The liquids content of the precipitated titanium hydroxide was gradually decreased as it passed through the filter 670 and the dewatering centrifuge 651, resulting in a final solid residue 655 of 58% by mass solids exiting the dewatering centrifuge 651 at a rate of 47 tpd (i.e. 27 solid tpd and 19 liquid tpd). At the same time, 2947 tpd of liquid (observed to be substantially free of Ti) 902 was recovered. The liquid 902 was noted to be a neutralised solution.

[473] Because the final solid residue 655 still comprised liquids (i.e. the remaining 42% by mass), it was transported to a rotary dryer for further drying. The rotary dryer was indirectly heated using combustion gases from the combustion of natural gas and was operated at a temperature above the decomposition temperature of titanium hydroxide. Advantageously, this was observed to result in a dried titanium dioxide product being produced, which could be directly sold as a valuable by-product. The dried titanium dioxide was produced at a rate of 27 tpd, i.e. all of the titanium dioxide that was contained in the impounded coal ash was recovered.

[474] As the titanium hydroxide was dried and decomposed, water was produced which evaporated and formed the vapour stream exiting the dryer at the discharge breech. The vapour stream was cooled to condense water vapour present, with the condensed water returned to the pond 102 at a rate of 19 tpd.

[475] The inventor noted that the neutralised solution 902 primarily comprised sodium nitrate and so could be used to produce a saline solution and regenerate the concentrated nitric acid 656. The solution 902 was pumped to a chlorination vessel 904 in which hydrochloric acid with a concentration of 10% was added. Sufficient hydrochloric acid was added to neutralise sodium hydroxide remaining in solution 902, i.e. producing sodium chloride and water. The aqueous solution 908 pumped from the chlorination vessel 904 was observed to comprise sodium cations, hydronium cations, chloride anions and nitrate anions. [476] The aqueous solution 908 was pumped to a heating vessel 910. The heating vessel 910 comprised a jacket through which steam was pumped, i.e. so as to provide heat to the vessel 910. The flow of steam was controlled such that a temperature of 85 °C was maintained inside the heating vessel 910. The inventor noted the flow of steam could be controlled manually or through process automation. At 85 °C, hydronium cations and nitrate anions in the solution 908 evaporated as a nitric acid vapour 912. The inventor noted that only a negligible amount of water was entrained with the nitric acid vapour 912 as it evaporated at 85 °C. The nitric acid vapour 912 evaporated at a rate of 1621 tpd.

[477] To recover the nitric acid, the vapour 912 was condensed 914 using a heat exchanger. Condensed nitric acid 916 was collected at a rate of 1621 tpd as essentially pure (100%) nitric acid. To regenerate the concentrated (68%) nitric acid 656, the 1621 tpd of condensed nitric acid 916 was combined with 763 tpd of recycled process water 920 in a dilution tank 918.

[478] The remaining ash sludge (now substantially free of Ti), separated as a slurry 660 by the centrifuge bank 658, was passed into the dilute-alkali digester 682. The dilute-alkali- digester 682 was a continuously stirred tank reactor. At the same time, a 1M sodium hydroxide solution 684 was added thereto, so as to promote the dissolution of Mo from the ash sludge as sodium molybdate. Sodium hydroxide 684 was added at a rate of 2345 tpd, as this was observed to be sufficient to recover substantially all the Mo from the ash sludge.

[479] To enable the separation of the resultant solution comprising Mo, the slurry 686 from the dilute-alkali digester 682 was passed to a centrifuge bank 688. A solution 692 comprising Mo was separated at a rate of 2345 tpd. At the same time, the remaining residual sludge was recovered as a slurry 690 comprising 50% solids by mass at a total rate of 4690 tpd, i.e. 2345 solid tpd and 2345 liquid tpd.

[480] The solution 692 comprising Mo was passed to an ion-exchange column which operated as a batch process. The ion-exchange column was packed with a resin which preferentially reacted with Mo. A concentrated Mo eluate 698 was produced at a rate of 0.5 tpd, which included 0.02 tpd of Mo (as oxide). That is, all the Mo that was in the impounded coal ash fed to the process 10 was recovered. The concentrated eluate 698 was dried to produce a dried Mo oxide product. The dried Mo oxide product was then added to the high- performance exotic metal alloy mix from the magnetic separation stage.

[481] A solution 696 substantially free of Mo was produced during the loading of the ionexchange column at an (average) rate of 2345 tpd. The solution 696 was typically used to produce the sodium hydroxide solution 684. However, the inventor noted that the solution 696 could also be recycled back to the pond 102.

[482] The remaining residual sludge recovered as a slurry 690 from the centrifuge bank 688 was added to the concentrated-alkali-digester 601. The concentrated-alkali-digester 601 was a continuously stirred tank reactor. At the same time, 18M sodium hydroxide 603 was added to the digester 601 so as to cause Ge in the residual sludge to dissolve therefrom.

The sodium hydroxide 603 was added at a rate of 2345 tpd, so as to ensure complete dissolution of Ge.

[483] To enable the separation of the resultant solution comprising Ge, the slurry 605 from the concentrated-alkali digester 601 was passed to a centrifuge bank 607. A solution 611 comprising Ge was separated at a rate of 2345 tpd. At the same time, the remaining residual sludge was recovered as a slurry 609 comprising 50% solids by mass at a total rate of 4690 tpd, i.e. 2345 solid tpd and 2345 liquid tpd.

[484] The solution 611 comprising Ge was added to a precipitation tank 613. Hydrochloric acid 615 at a concentration of 35% by mass was also added to the precipitation tank 613, at a rate of 2751 tpd. This caused the Ge in solution to precipitate as germanium hydroxide.

[485] To enable recovery of the precipitated germanium hydroxide, the resultant slurry 617 from the precipitation tank 613 was passed through a series of solid/liquid separation stages including a filter 619 and a dewatering centrifuge 657. Spray water was added to the filter 619 at a rate of 1 tpd to help rinse the precipitated germanium hydroxide.

[486] The liquids content of the precipitated germanium hydroxide was gradually decreased as it passed through the filter 619 and the dewatering centrifuge 657, resulting in a final solid residue 661 of 6% by mass solids exiting the dewatering centrifuge 657 at a rate of 0.69 tpd (i.e. 0.04 solid tpd and 0.65 liquid tpd). At the same time, 5096 tpd of liquid (observed to be substantially free of Ge) 623, 659 was recovered. The liquid 623, 659 was observed to be saline and so was suitable for use in the chlor-alkali plant or for the regeneration of resin in ion-exchange columns.

[487] Because the final solid residue 661 still comprised liquids (i.e. the remaining 94% by mass), it was transported to a rotary dryer for further drying. The rotary dryer was indirectly heated using combustion gases from the combustion of natural gas and was operated at a temperature above the decomposition temperature of germanium hydroxide.

Advantageously, this was observed to result in a dried germanium dioxide product being produced, which could be sold as a valuable by-product. The dried germanium dioxide was produced at a rate of 0.04 tpd, i.e. all of the germanium dioxide that was contained in the impounded coal ash was recovered.

[488] As the germanium hydroxide was dried and decomposed, water was produced which evaporated and formed part of the vapour stream exiting the dryer at the discharge breech. The vapour stream was cooled to condense water vapour present, with the condensed water returned to the pond 627 at a rate of 0.65 tpd.

[489] The inventor noted that the solid component of the remaining residual sludge recovered as a slurry 609 from the centrifuge bank 688 was an inert mixture of silica and zirconia. Hydrochloric acid 665 was added to the slurry 609 until the residual sludge had a pH of between 7 and 8.

[490] To enable recovery of the inert silica/zirconia solid, the (neutralised) slurry 609 was passed to a dewatering screw press conveyor 631. As the solid was squeezed in the screw press conveyor 631, liquids were removed at a rate of 938 tpd. The recovered liquid 635 was observed to be alkaline and was used to generate the sodium hydroxide solutions required elsewhere in the process (e.g. 684, 603).

[491] By squeezing out the liquids, the screw press 631 also produced a compacted mass of solid 633 comprising silica/zirconia which exited the upper end of the screw press at a rate of 3752 tpd, consisting of 2345 solid tpd and 1407 liquid tpd. That is, the moisture content of the solid 633 exiting the screw press was reduced to -38% by mass.

[492] The inventor noted that the solid 633 was suitable for sale and use as a lightweight manufactured earth fill 637 in road construction and for mine void fillings. However, typically, the solid 633 was passed to an indirectly heated rotary dryer 641 for further drying.

[493] In the rotary dryer 641, the solid 639 was dried, thereby producing a solid dried product 643 which was free of liquids at a rate of 2345 tpd. That is, all the silica and zirconia that entered the process 10 in the impounded coal ash was recovered in the solid dried product 643. The inventor noted that the solid dried product 643 comprised 2343 tpd of silica and 1 tpd of zirconia and was suitable for use as a component of a cementitious product, such as the structural light weight aggregate product of AU 2021106969.

[494] Water present in the solid 639 was evaporated as the solid was dried and exited the dryer 641 as a vapour. The vapour was cooled to condense the water 645, which was recovered at a rate of 1407 tpd. Out of the recovered water, 763 tpd was recycled to make up the concentrated nitric acid 656 and the remaining 644 tpd was pumped back to the pond 102. Acid-Soluble Recovery Area

[495] The solution 502 comprising the acid-soluble components was pumped to the acidsoluble recovery area 700 at a rate of 4219 tpd. It was observed that the solution 502 comprised the acid-soluble components Sb, Be, Cd, Zn, Sn, Al, Ca and Mg.

[496] Fig. 7 shows the stages of the acid-soluble recovery area 700. In particular, the acidsoluble recovery area 700 comprised five ion-exchange stages 703, 708, 714, 720, 726. Each ion-exchange stage was aimed at removing one of Sb, Be, Cd, Zn, Sn from the solution 502, i.e. such that a solution comprising only Al, Ca and Mg remained. Suitable resins were selected for each ion-exchange stage 703, 708, 714, 720, 726 to enable the components to be recovered in the order Sb, Be, Cd, Zn, Sn respectively.

[497] Each stage 703, 708, 714, 720, 726 comprised a single column and was operated as a batch process, i.e. the column was first loaded to remove the ion of interest and was then eluted to generate the concentrated eluate and regenerate the resin. The concentrated eluates of Sb, Be, Cd, Zn, Sn were pumped into storage tanks 732, 734, 736, 738, 740 respectively for storage.

[498] The inventor noted that substantially all of the Sb, Be, Cd, Zn, Sn present in the coal ash dredged from the pond 102 was able to be recovered in the acid-soluble recovery area. In particular, 0.01 tpd of antinomy oxide, 0.05 tpb of beryllium oxide, 0.001 tpd of cadmium oxide, 0.23 tpd of zinc oxide and 0.02 tpd of tin oxide were recovered into storage tanks 732, 734, 736, 738, 740 respectively. The concentrated eluates of Sb, Be, Cd, Sn were also observed to comprise 0.5 tpd of liquids, whilst the concentrated eluate of Zn comprised 1.0 tpd of liquids.

[499] The inventor noted that all of the components Sb, Be, Cd, Zn, Sn were suitable for use in a structural light weight aggregate, such as the type described in 2021902692.

Because the quantities of concentrated eluates of Sb, Be, Cd, Zn, Sn were small (i.e. all less than 1.0 tpd) relative to the size of the storage tanks 732, 734, 736, 738, 740, the storage tanks were only periodically emptied (i.e. as they filled up). When the storage tanks 732, 734, 736, 738, 740 needed to be emptied, the contents were transported to the dryer located prior to the particle separation stage of AU 2021902692.

[500] From the final ion-exchange stage 726, a solution 728 comprising Al, Ca and Mg was recovered. The inventor noted that the Al, Ca and Mg were in the form of dissolved chlorides. The solution 704 had a flow rate of 4219 tpd, which included 923 tpd of dissolved solids (i.e. the chlorides of Al, Ca and Mg). That is, the flow rate of the solution 704 was substantially the same as the mass flow rate of the solution 502 entering the first ionexchange stage 703 because the mass of the components recovered in the five ionexchange stages 703, 708, 714, 720, 726 was small.

Precipitation Area

[501] The solution 728 comprising Al, Ca and Mg was pumped to a precipitation tank 803 in the precipitation area 800 (Fig. 8). The precipitation tank 803 was a continuously stirred tank reactor, i.e. the precipitation area 800 was a continuous process. However, the inventor noted that the precipitation area 800 could also be operated as a series of batch processes.

[502] Sodium carbonate 805 was added to the precipitation tank 803 so as to promote the conversion of the Al, Ca and Mg chlorides to alumina, calcium carbonate and magnesium carbonate respectively. The inventor noted that as the process 10 was starting up, sodium carbonate 805 was added as (solid) sodium carbonate decahydrate crystals. However, when the process was operating at a steady-state, only a minimal quantity of fresh sodium carbonate was required. This was because some sodium carbonate was regenerated from carbon dioxide released by the formation of alumina and from the reactions that occurred in the dissolution vessel 810 downstream.

[503] The inventor also noted that the conditions of the precipitation tank 803 were carefully controlled so as to avoid the formation of calcium aluminate, because this would detrimentally affect the purity of the alumina product.

[504] The inventor noted that exiting the precipitation tank 803 was a slurry 808 comprising a precipitate of an alumina, calcium carbonate and magnesium carbonate. The slurry 808 was pumped to dissolution vessel 810, which was also a continuously stirred tank reactor.

[505] Cool hydrochloric acid 812 at a concentration of 35% by mass was added to the dissolution vessel 810 at a rate of 923 tpd. The hydrochloric acid was cool so as to promote the dissolution of calcium carbonate and magnesium carbonate as chlorides from the precipitate, without causing the alumina to dissolve therefrom. Thus, the remaining solid substantially comprised alumina.

[506] The carbon dioxide generated as a result of the dissolution of calcium carbonate and magnesium carbonate was captured and bubbled through a solution of sodium hydroxide, thereby regenerating sodium carbonate. The regenerated sodium carbonate was added to the precipitation tank 802, thus reducing the requirement for fresh sodium carbonate. [507] The slurry 816 from the dissolution vessel 810 was pumped to a centrifuge bank 818, enabling the solid substantially comprising alumina to be separated. The solid substantially comprising alumina was separated as a slurry 820 comprising 50% by mass solids, with a total mass flow of 1694 tpd (i.e. 847 solid tpd and 847 liquid tpd). At the same time, the liquids 822 separated from the solid substantially comprising alumina were recovered at a rate of 4296 tpd. The liquids 822 comprised 76 tpd of dissolved solids, i.e. the liquids 822 was a solution comprising the redissolved Ca and Mg chlorides.

[508] Because the solid substantially comprising alumina was separated as a slurry comprising liquids (i.e. the remaining 50% by mass), the slurry 820 was transported to a screw press 824 for dewatering. In the screw press 824, the liquids were squeezed out of the solid substantially comprising alumina and were collected at a centrate outlet thereof at a rate of 339 tpd. Because the collected liquids 828 were slightly acidic, they were recycled to the non-acid-soluble recovery area.

[509] By squeezing out the liquids, the screw press also produced a compacted mass of solid substantially comprising alumina 826 which exited the upper end of the screw press at a rate of 1355 tpd, consisting of 847 solid tpd and 508 liquid tpd. That is, the moisture content of the solid substantially comprising alumina 826 exiting the screw press was reduced to -38% by mass.

[510] From the outlet of the screw press, the solid substantially comprising alumina 826 was directed to an indirectly-heated rotary kiln dryer 830 for further drying. In the dryer 830, the solid substantially comprising alumina was dried as it contacted the internal walls and vanes of the dryer 830 which were heated due to the combustion of natural gas. The dried alumina solid 802 exiting the dryer 830 was then sold. The inventor observed that the dried alumina solid 802 was produced at a rate of 847 tpd. This was equivalent to the rate at which alumina entered the process 10 in the dredged coal ash from the pond 102, i.e. all the alumina that entered the process 10 was recovered.

[511] As the solid 802 substantially comprising alumina was dried, water evaporated from the solid entered the vapour phase. Thus, the vapour 838 exiting the dryer comprised 508 tpd water. The vapour 838 was cooled to condense the water, which was pumped back to the pond 102.

[512] The solution 822 comprising the redissolved Ca and Mg chlorides was pumped to a second precipitation tank 840, which was likewise a continuously stirred tank reactor.

Sodium carbonate was added to the precipitation tank 840 so as to neutralise the hydrochloric acid and reprecipitate the Ca and Mg as calcium carbonate and magnesium carbonate respectively.

[513] As the hydrochloric acid was neutralised, carbon dioxide 862 was produced. The carbon dioxide 862 was captured and bubbled through a solution comprising sodium hydroxide, thereby regenerating a solution comprising sodium carbonate. Said solution was recycled back to the precipitation tank 840 for re-use.

[514] The inventor again noted that, during start-up, the sodium carbonate added to the precipitation tank 840 in the form of sodium carbonate decahydrate crystals was the only source of sodium carbonate available. However, as the process reached a steady-state, the solution comprising sodium carbonate was regenerated via the captured carbon dioxide and was added back to the precipitation tank 840. This decreased the requirement for fresh sodium carbonate.

[515] From the precipitation tank 840, a slurry comprising a calcium carbonate and magnesium carbonate precipitate and a solution substantially free of Ca and Mg was pumped to a centrifuge bank 846. The centrifuge bank 846 enabled the solid comprising the calcium carbonate and magnesium carbonate precipitate to be separated from the solution. It was observed that the calcium carbonate and magnesium carbonate were in the form of dolomite and calcite.

[516] The solid comprising dolomite and calcite was separated as a slurry 848 comprising 50% by mass solids, with a total mass flow of 153 tpd (i.e. 76 solid tpd and 76 liquid tpd). At the same time, the liquids 850 separated from the solid substantially comprising alumina were recovered at a rate of 4326 tpd. The liquids 850 were observed to be substantially free of Al, Ca and Mg, but did comprise sodium chloride. Because the liquids 850 were saline, they were suitable for reuse in either the chlor-alkali plant or for use in the regeneration of ion-exchange resins.

[517] Because the solid 848 comprising dolomite and calcite was separated as a slurry (i.e. comprising 50% by mass liquids), the slurry 848 was transported to a screw press 864 for dewatering. In the screw press 864, the liquids were squeezed out of the solid comprising dolomite and calcite and were collected at a centrate outlet thereof at a rate of 31 tpd.

Because the collected liquids 854 were saline, they were combined with the liquids 850 from the centrifuge bank 846.

[518] By squeezing out the liquids, the screw press 864 also produced a compacted mass of solid 852 comprising dolomite and calcite which exited the upper end of the screw press at a rate of 122 tpd, consisting of 76 solid tpd and 46 liquid tpd. That is, the moisture content of the solid 852 comprising dolomite and calcite exiting the screw press was reduced to -29% by mass.

[519] From the outlet of the screw press, the solid 852 comprising dolomite and calcite was directed to an indirectly-heated rotary kiln dryer 856 for further drying. In the dryer 856, the solid comprising dolomite and calcite 852 was dried as it contacted the internal walls and vanes of the dryer 856 which were heated from the combustion of natural gas in the burners. A dried dolomite and calcite solid product 804 exited the dryer 830. The inventor observed that the mass of the dried dolomite and calcite solid product 804 of 76 solid tpd corresponded to the mass of Ca (as hydroxide) and Mg (as oxide) entering the process 10 in the coal ash dredged from the pond 102, i.e. all the Ca and Mg was recovered in the product 804.

[520] As the solid 852 comprising dolomite and calcite was dried, water evaporated from the solid entered the vapour phase. Thus, the vapour 860 exiting the dryer comprised 46 tpd water. The vapour 860 was cooled to condense the water, which was pumped back to the pond 102.

[521] The inventor noted that, depending on the concentration of Mg and Ca in the dried solid 804 comprising dolomite and calcite, the final product could have different end-uses. For example, the solid 804 could be further treated to separate a high magnesium component from a low magnesium component, each of which could be sold as individual products. Typically, however, the inventor noted that the solid 804 was used as a component of a structural light weight aggregate.

[522] A chlor-alkali plant 950 was also present as part of the process 10. Solutions comprising sodium chloride generated by the process 10 were used as feed 952 to the chloralkali plant 950 in the first compartment 954 of the cell 956. The inventor noted that the solution 952 was already at a concentration of -26% NaCI by mass. However, the chloralkali plant 950 comprised a reverse osmosis unit which could be used to further concentrate the solutions comprising sodium chloride, when the concentration was below 26% NaCI by mass. Fresh water 961 was fed to the second compartment 957 of the cell 956. The inventor noted that the fresh water 961 could comprise recovered process water, e.g. condensate from the dryers.

[523] The chlor-alkali plant produced hydrogen gas 966 and chlorine gas 964 which were used to generate hydrochloric acid for use in the process 10. [524] The chlor-alkali plant also produced a solution comprising sodium hydroxide 968 which was recovered for use in the process 10. The inventor noted that the sodium hydroxide solution 968 could be further concentrated (e.g. by evaporation) as required.

Further Variations

[525] Variations and modifications may be made to the process as previously described without departing from the spirit or ambit of the disclosure.

[526] For example, it is to be understood that the characteristics of coal ash may differ to the extent that variations to the above process may be appropriate. Other unit operations can be included in the overall process in line with good engineering practices, in particular, for the extraction and recovery of metals, the conservation of water, and the minimisation of waste streams.

[527] 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” or variations such as “comprises” or “comprising” is 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.