Processes for preparing alumina and magnesium chloride by HCl leaching of various materials

Abstract
The disclosed processes can be effective for treating various materials comprising several different metals. These materials can be leached with HCl for obtaining a leachate and a solid. Then, they can be separated from one another and a first metal can be isolated from the leachate. Then, a second metal can further be isolated from the leachate. The first and second metals can each be substantially selectively isolated from the leachate. This can be done by controlling the temperature of the leachate, adjusting pH, further reacting the leachate with HCl, etc. The metals that can be recovered in the form of metal chlorides can eventually be converted into the corresponding metal oxides, thereby allowing for recovering HCl. The various metals can be chosen from aluminum, iron, zinc, copper, gold, silver, molybdenum, cobalt, magnesium, lithium, manganese, nickel, palladium, platinum, thorium, phosphorus, uranium, titanium, rare earth element and rare metals.
Description
TECHNICAL FIELD

The present disclosure relates to improvements in the field of chemistry applied to the treatment of various ores. For example, it relates to processes for treating materials comprising at least one metal chosen from aluminum, iron, zinc, copper, gold, silver, molybdenum, cobalt, magnesium, lithium, manganese, nickel, palladium, platinum, thorium, phosphorus, uranium and titanium, and/or at least one rare earth element and/or at least one rare metal.


BACKGROUND OF THE DISCLOSURE

There have been several known processes for the production of alumina, titanium oxide, magnesium oxide, hematite, nickel, cobalt rare earth elements, rare metals etc. Many of them have the disadvantage of being inefficient to segregate and extract value added secondary products, thus leaving an important environmental impact.


SUMMARY OF THE DISCLOSURE

According to one aspect, there is provided a process for preparing alumina and optionally other products, the process comprising:

    • leaching an aluminum-containing material with HCl so as to obtain a leachate comprising aluminum ions and a solid, and separating the solid from the leachate;
    • reacting the leachate with HCl so as to obtain a liquid and a precipitate comprising the aluminum ions in the form of AlCl3, and separating the precipitate from the liquid;
    • heating the precipitate under conditions effective for converting AlCl3 into Al2O3 and recovering gaseous HCl so-produced; and
    • recycling the gaseous HCl so-produced by contacting it with water so as to obtain a composition having a concentration higher than HCl azeotrope concentration (20.2 weight %) and reacting the composition with a further quantity of aluminum-containing material so as to leaching it.


According to another aspect, there is provided a process for preparing alumina and optionally other products, the process comprising:

    • leaching an aluminum-containing material with HCl so as to obtain a leachate comprising aluminum ions and a solid, and separating the solid from the leachate;
    • reacting the leachate with HCl so as to obtain a liquid and a precipitate comprising the aluminum ions in the form of AlCl3, and separating the precipitate from the liquid; and
    • optionally reacting the precipitate with a base; and
    • heating the precipitate under conditions effective for converting it into Al2O3.


According to another aspect, there is provided a process for preparing alumina and optionally other products, the process comprising:

    • leaching an aluminum-containing material with HCl so as to obtain a leachate comprising aluminum ions and a solid, and separating the solid from the leachate;
    • reacting the leachate with HCl so as to obtain a liquid and a precipitate comprising the aluminum ions in the form of AlCl3, and separating the precipitate from the liquid; and
    • optionally reacting the precipitate with a base; and
    • heating the precipitate under conditions effective for converting it into Al2O3.


According to another aspect, there is provided a process for preparing alumina and optionally other products, the process comprising:

    • leaching an aluminum-containing material with HCl so as to obtain a leachate comprising aluminum ions and a solid, and separating the solid from the leachate;
    • reacting the leachate with HCl so as to obtain a liquid and a precipitate comprising the aluminum ions in the form of AlCl3, and separating the precipitate from the liquid;
    • heating the precipitate under conditions effective for converting AlCl3 into Al2O3 and recovering gaseous HCl so-produced; and
    • recycling the gaseous HCl so-produced by contacting it with water so as to obtain a composition having a concentration of about 18 to about 45 weight % or about 25 to about 45 weight % and reacting the composition with a further quantity of aluminum-containing material so as to leaching it.


According to another aspect, there is provided a process for preparing alumina and optionally other products, the process comprising:

    • leaching an aluminum-containing material with HCl so as to obtain a leachate comprising aluminum ions and a solid, and separating the solid from the leachate;
    • reacting the leachate with HCl so as to obtain a liquid and a precipitate comprising the aluminum ions in the form of AlCl3, and separating the precipitate from the liquid;
    • heating the precipitate under conditions effective for converting AlCl3 into Al2O3 and recovering gaseous HCl so-produced; and
    • recycling the gaseous HCl so-produced by contacting it with water so as to obtain a composition having a concentration of about 18 to about 45 weight % or about 25 to about 45 weight % and using the composition for leaching the aluminum-containing material.


According to another aspect, there is provided a process for preparing alumina and optionally other products, the process comprising:

    • leaching an aluminum-containing material with HCl so as to obtain a leachate comprising aluminum ions and a solid, and separating the solid from the leachate;
    • reacting the leachate with HCl so as to obtain a liquid and a precipitate comprising the aluminum ions in the form of AlCl3, and separating the precipitate from the liquid;
    • heating the precipitate under conditions effective for converting AlCl3 into Al2O3 and recovering gaseous HCl so-produced; and
    • recycling the gaseous HCl so-produced by contacting it with the leachate so as to precipitate the aluminum ions in the form of AlCl3.6H2O.


According to another aspect, there is provided a process for preparing alumina and optionally other products, the process comprising:

    • leaching an aluminum-containing material with HCl so as to obtain a leachate comprising aluminum ions and a solid, and separating the solid from the leachate;
    • reacting the leachate with HCl so as to obtain a liquid and a precipitate comprising the aluminum ions in the form of AlCl3, and separating the precipitate from the liquid; and
    • heating the precipitate under conditions effective for converting AlCl3 into Al2O3.


According to another aspect, there is provided a process for preparing alumina and optionally other products, the process comprising:

    • leaching an aluminum-containing material with HCl so as to obtain a leachate comprising aluminum ions and a solid, and separating the solid from the leachate;
    • reacting the leachate with HCl so as to obtain a liquid and a precipitate comprising the aluminum ions in the form of AlCl3, and separating the precipitate from the liquid; and
    • heating the precipitate under conditions effective for converting AlCl3 into Al2O3 and optionally recovering gaseous HCl so-produced.


According to one aspect, there is provided a process for preparing aluminum and optionally other products, the process comprising:

    • leaching an aluminum-containing material with HCl so as to obtain a leachate comprising aluminum ions and a solid, and separating the solid from the leachate;
    • reacting the leachate with HCl so as to obtain a liquid and a precipitate comprising the aluminum ions in the form of AlCl3, and separating the precipitate from the liquid;
    • heating the precipitate under conditions effective for converting AlCl3 into Al2O3; and
    • converting Al2O3 into aluminum.


According to another aspect, there is provided a process for preparing aluminum and optionally other products, the process comprising:

    • leaching an aluminum-containing material with HCl so as to obtain a leachate comprising aluminum ions and a solid, and separating the solid from the leachate;
    • reacting the leachate with HCl so as to obtain a liquid and a precipitate comprising the aluminum ions in the form of AlCl3, and separating the precipitate from the liquid;
    • heating the precipitate under conditions effective for converting AlCl3 into Al2O3 and optionally recovering gaseous HCl so-produced; and
    • converting Al2O3 into aluminum.


According to another aspect, there is provided a process for preparing various products, the process comprising:

    • leaching a material comprising a first metal with HCl so as to obtain a leachate comprising ions of the first metal and a solid, and separating the solid from the leachate;
    • reacting the leachate with HCl so as to obtain a liquid and a precipitate comprising a chloride of the first metal, and separating the precipitate from the liquid; and
    • heating the precipitate under conditions effective for converting the chloride of the first metal into an oxide of the first metal.


According to another aspect, there is provided a process for treating serpentine, the process comprising:

    • leaching serpentine with HCl so as to obtain a leachate comprising magnesium ions and a solid, and separating the solid from the leachate;
    • reacting the leachate with HCl so as to obtain a liquid and a precipitate comprising MgCl2, and separating the precipitate from the liquid; and
    • heating the precipitate under conditions effective for converting MgCl2 into MgO and optionally recovering gaseous HCl so-produced.


According to another aspect, there is provided a process for treating serpentine, the process comprising:

    • leaching serpentine with HCl so as to obtain a leachate comprising magnesium ions and a solid, and separating the solid from the leachate;
    • reacting the leachate with HCl so as to obtain a liquid and a precipitate comprising MgCl2, and separating the precipitate from the liquid; and
    • heating the precipitate under conditions effective for converting MgCl2 into MgO.


According to another aspect, there is provided process for treating a magnesium-containing material, the process comprising:

    • leaching the magnesium-containing material with HCl so as to obtain a leachate comprising magnesium ions and a solid, and separating the solid from the leachate;
    • reacting the leachate with HCl so as to obtain a liquid and a precipitate comprising MgCl2, and separating the precipitate from the liquid; and
    • heating the precipitate under conditions effective for converting MgCl2 into MgO and optionally recovering gaseous HCl so-produced.


According to another aspect, there is provided a process for treating a magnesium-containing material, the process comprising:

    • leaching the magnesium-containing material with HCl so as to obtain a leachate comprising magnesium ions and ions from at least one metal and a solid, and separating the solid from the leachate; and
    • precipitating the at least one metal by reacting the leachate with a precipitating agent so as to obtain a liquid comprising the magnesium ions and a precipitate comprising the precipitated at least one metal, and separating the precipitate from the liquid.


According to another aspect, there is provided a process for treating a material comprising magnesium and at least one other metal, the process comprising:

    • leaching the material with HCl so as to obtain a leachate comprising magnesium ions and ions from the at least one other metal and a solid, and separating the solid from the leachate; and
    • precipitating the at least one other metal by reacting the leachate with a precipitating agent so as to obtain a liquid comprising the magnesium ions and a precipitate comprising the precipitated at least one metal, and separating the precipitate from the liquid;
    • treating the liquid so as to cause precipitation of Mg(OH)2; and
    • treating the precipitate so as to substantially selectively isolate the at least one metal therefrom.


According to another aspect, there is provided a process for preparing alumina, the process comprising:

    • leaching an aluminum-containing material with HCl so as to obtain a leachate comprising aluminum ions, magnesium ions and a solid, and separating the solid from the leachate;
    • substantially selectively precipitating MgCl2 from the leachate and removing the MgCl2 from the leachate;
    • reacting the leachate with HCl so as to obtain a liquid and a precipitate comprising the aluminum ions in the form of AlCl3, and separating the precipitate from the liquid;
    • heating the precipitate under conditions effective for converting AlCl3 into Al2O3 and optionally recovering gaseous HCl so-produced; and
    • heating the MgCl2 under conditions effective for converting it into MgO and optionally recovering gaseous HCl so-produced.


According to another aspect, there is provided a process for preparing aluminum, the process comprising:

    • leaching an aluminum-containing material with HCl so as to obtain a leachate comprising aluminum ions, magnesium ions and a solid, and separating the solid from the leachate;
    • substantially selectively precipitating MgCl2 from the leachate and removing the MgCl2 from the leachate;
    • reacting the leachate with HCl so as to obtain a liquid and a precipitate comprising the aluminum ions in the form of AlCl3, and separating the precipitate from the liquid;
    • heating the MgCl2 under conditions effective for converting it into MgO and optionally recovering gaseous HCl so-produced;
    • heating the precipitate under conditions effective for converting AlCl3 into Al2O3 and optionally recovering gaseous HCl so-produced; and
    • converting the Al2O3 into alumina.


According to another aspect, there is provided a process for preparing alumina, the process comprising:

    • leaching an aluminum-containing material with HCl so as to obtain a leachate comprising aluminum ions, magnesium ions and a solid, and separating the solid from the leachate;
    • substantially selectively precipitating MgCl2 from the leachate and removing the MgCl2 from the leachate;
    • reacting the leachate with HCl so as to obtain a liquid and a precipitate comprising the aluminum ions in the form of AlCl3, and separating the precipitate from the liquid;
    • optionally treating the precipitate with a base;
    • heating the precipitate under conditions effective for converting the precipitate into Al2O3 and optionally recovering gaseous HCl so-produced; and
    • heating the MgCl2 under conditions effective for converting it into MgO and optionally recovering gaseous HCl so-produced.


According to another aspect, there is provided a process for preparing aluminum, the process comprising:

    • leaching an aluminum-containing material with HCl so as to obtain a leachate comprising aluminum ions, magnesium ions and a solid, and separating the solid from the leachate;
    • substantially selectively precipitating MgCl2 from the leachate and removing the MgCl2 from the leachate;
    • reacting the leachate with HCl so as to obtain a liquid and a precipitate comprising the aluminum ions in the form of AlCl3, and separating the precipitate from the liquid;
    • optionally treating the precipitate with a base;
    • heating the MgCl2 under conditions effective for converting it into MgO and optionally recovering gaseous HCl so-produced;
    • heating the precipitate under conditions effective for converting the precipitate into Al2O3 and optionally recovering gaseous HCl so-produced; and
    • converting the Al2O3 into alumina.


According to another aspect, there is provided a process for treating serpentine, the process comprising:

    • leaching serpentine with HCl so as to obtain a leachate comprising magnesium ions and a solid, and separating the solid from the leachate;
    • controlling the temperature of the leachate so as to substantially selectively precipitate the magnesium ions in the form of magnesium chloride, and removing the precipitate from the leachate, thereby obtaining a liquid; and
    • heating the MgCl2 under conditions effective for converting MgCl2 into MgO and optionally recovering gaseous HCl so-produced.


According to another aspect, there is provided a process for treating a magnesium-containing material, the process comprising:

    • leaching the magnesium-containing material with HCl so as to obtain a leachate comprising magnesium ions, and a solid, and separating the solid from the leachate;
    • controlling the temperature of the leachate so as to substantially selectively precipitate the magnesium ions in the form of magnesium chloride, and removing the precipitate from the leachate, thereby obtaining a liquid; and
    • heating the MgCl2 under conditions effective for converting MgCl2 into MgO and optionally recovering gaseous HCl so-produced.


According to another aspect, there is provided a process for preparing various products, the process comprising:

    • leaching a material comprising a first metal with HCl so as to obtain a leachate comprising ions of the first metal and a solid, and separating the solid from the leachate;
    • controlling the temperature of the leachate so as to precipitate the first metal in the form of a chloride, and removing the precipitate from the leachate,
    • reacting the leachate with HCl so as to obtain a liquid and a precipitate comprising a chloride of the second metal, and separating the precipitate from the liquid;
    • optionally heating the chloride of the first metal under conditions effective for converting it into an oxide of the first metal, and optionally recovering the so-produced HCl; and
    • optionally heating the chloride of the second metal under conditions effective for converting it into an oxide of the second metal, and optionally recovering the so-produced HCl.


According to another aspect, there is provided a process for preparing various products, the process comprising:

    • leaching a material comprising a first metal with HCl so as to obtain a leachate comprising ions of the first metal and a solid, and separating the solid from the leachate;
    • controlling the temperature of the leachate so as to precipitate the first metal in the form of a chloride, and removing the precipitate from the leachate;
    • controlling the temperature of the leachate so as to precipitate the second metal in the form of a chloride, and removing the precipitate from the leachate;
    • optionally heating the chloride of the first metal under conditions effective for converting it into an oxide of the first metal, and optionally recovering the so-produced HCl; and
    • optionally heating the chloride of the second metal under conditions effective for converting it into an oxide of the second metal, and optionally recovering the so-produced HCl.


According to another aspect, there is provided a process for preparing various products, the process comprising:

    • leaching a material comprising a first metal with HCl so as to obtain a leachate comprising ions of the first metal and a solid, and separating the solid from the leachate;
    • reacting the leachate with HCl so as to obtain a precipitate comprising the first metal in the form of a chloride, and removing the precipitate from the leachate;
    • reacting the leachate with HCl so as to obtain a precipitate comprising a second metal in the form of a chloride, and removing the precipitate from the leachate;
    • optionally heating the chloride of the first metal under conditions effective for converting it into an oxide of the first metal, and optionally recovering the so-produced HCl; and
    • optionally heating the chloride of the second metal under conditions effective for converting it into an oxide of the second metal, and optionally recovering the so-produced HCl.


According to another aspect, there is provided a process for preparing various products, the process comprising:

    • leaching a material comprising a first metal with HCl so as to obtain a leachate comprising ions of the first metal and a solid, and separating the solid from the leachate;
    • controlling the concentration of HCl in the leachate and/or the temperature of the leachate so as to precipitate the first metal in the form of a chloride, and removing the precipitate from the leachate;
    • controlling the concentration of HCl in the leachate and/or the temperature of the leachate so as to precipitate a second metal in the form of a chloride, and removing the precipitate from the leachate;
    • optionally heating the chloride of the first metal under conditions effective for converting it into an oxide of the first metal, and optionally recovering the so-produced HCl; and
    • optionally heating the chloride of the second metal under conditions effective for converting it into an oxide of the second metal, and optionally recovering the so-produced HCl.


According to another aspect, there is provided a process for preparing various products, the process comprising:

    • leaching a material comprising a first metal with HCl so as to obtain a leachate comprising ions of the first metal and a solid, and separating the solid from the leachate;
    • controlling the concentration of HCl in the leachate and/or the temperature of the leachate so as to precipitate the first metal in the form of a chloride, and removing the precipitate from the leachate;
    • reacting the leachate with HCl so as to obtain a precipitate comprising a second metal in the form of a chloride, and removing the precipitate from the leachate,
    • optionally heating the chloride of the first metal under conditions effective for converting it into an oxide of the first metal, and optionally recovering the so-produced HCl; and
    • optionally heating the chloride of the second metal under conditions effective for converting it into an oxide of the second metal, and optionally recovering the so-produced HCl.


According to another aspect, there is provided a process for preparing various products, the process comprising:

    • leaching a material comprising a first metal with HCl so as to obtain a leachate comprising ions of the first metal and a solid, and separating the solid from the leachate;
    • reacting the leachate with HCl so as to obtain a precipitate comprising the first metal in the form of a chloride, and removing the precipitate from the leachate;
    • controlling the concentration of HCl in the leachate and/or the temperature of the leachate so as to precipitate a second metal in the form of a chloride, and removing the precipitate from the leachate;
    • optionally heating the chloride of the first metal under conditions effective for converting it into an oxide of the first metal, and optionally recovering the so-produced HCl; and
    • optionally heating the chloride of the second metal under conditions effective for converting it into an oxide of the second metal, and optionally recovering the so-produced HCl.


According to another aspect, there is provided a process for preparing various products, the process comprising:

    • leaching a material comprising magnesium and iron with HCl so as to obtain a leachate comprising magnesium ions and iron ions and a solid, and separating the solid from the leachate;
    • reacting the leachate with HCl so as to obtain a precipitate comprising magnesium chloride, and removing the precipitate from the leachate so as to obtain a liquid comprising iron chloride;
    • treating the liquid under conditions effective for converting the iron chloride into iron oxide and optionally recovering HCl; and
    • optionally heating the magnesium chloride under conditions effective for converting it into magnesium oxide, and optionally recovering the so-produced HCl.


According to another aspect, there is provided a process for preparing various products, the process comprising:

    • leaching a material comprising magnesium and iron with HCl so as to obtain a leachate comprising magnesium ions and iron ions and a solid, and separating the solid from the leachate;
    • controlling the concentration of HCl in the leachate and/or the temperature of the leachate so as to precipitate magnesium chloride, and removing the precipitate from the leachate, thereby obtaining a liquid;
    • treating the liquid under conditions effective for converting the iron chloride into iron oxide and optionally recovering HCl; and
    • optionally heating the magnesium chloride under conditions effective for converting it into magnesium oxide, and optionally recovering HCl.


According to another aspect, there is provided a process for preparing various products, the process comprising:

    • leaching a material comprising magnesium, aluminum and iron with HCl so as to obtain a leachate comprising magnesium ions, aluminum ions and iron ions and a solid, and separating the solid from the leachate;
    • controlling the concentration of HCl in the leachate and/or the temperature of the leachate so as to precipitate magnesium chloride, and removing the precipitate from the leachate,
    • reacting the leachate with HCl so as to obtain a precipitate comprising aluminum chloride, and removing the precipitate from the leachate so as to obtain a liquid comprising iron chloride;
    • optionally treating the liquid under conditions effective for converting the iron chloride into iron oxide and optionally recovering HCl;
    • optionally heating the precipitate under conditions effective for converting aluminum chloride into alumina and optionally recovering gaseous HCl so-produced; and
    • optionally heating the magnesium chloride under conditions effective for converting it into magnesium oxide, and optionally recovering HCl.


According to another aspect, there is provided a process for preparing various products, the process comprising:

    • leaching a material comprising magnesium, aluminum and iron with HCl so as to obtain a leachate comprising magnesium ions, aluminum ions and iron ions and a solid, and separating the solid from the leachate;
    • reacting the leachate with HCl so as to obtain a precipitate comprising aluminum chloride, and removing the precipitate from the leachate;
    • controlling the concentration of HCl in the leachate and/or the temperature of the leachate so as to precipitate magnesium chloride, and removing the precipitate from the leachate so as to obtain a liquid comprising iron chloride
    • optionally treating the liquid under conditions effective for converting the iron chloride into iron oxide and optionally recovering HCl;
    • optionally heating the precipitate under conditions effective for converting aluminum chloride into alumina and optionally recovering gaseous HCl so-produced; and
    • optionally heating the magnesium chloride under conditions effective for converting it into magnesium oxide, and optionally recovering HCl.





BRIEF DESCRIPTION OF DRAWINGS

In the following drawings, which represent by way of example only, various embodiments of the disclosure:



FIG. 1 shows a bloc diagram of an example of process for preparing alumina and various other products according to the present disclosure;



FIG. 2 is an extraction curve for Al and Fe in which the extraction percentage is expressed as a function of a leaching time in a process according to an example of the present application;



FIG. 3 shows a bloc diagram of another example of process for preparing alumina and various other products according to the present disclosure;



FIG. 4 is a schematic representation of an example of a process for purifying/concentrating HCl according to the present disclosure;



FIG. 5 is a schematic representation of an example of a process for purifying/concentrating HCl according to the present disclosure;



FIG. 6 shows another bloc diagram of an example of process for preparing alumina and various other products according to the present disclosure;



FIG. 7 shows another bloc diagram of an example of process for preparing alumina and various other products according to the present disclosure;



FIG. 8 shows another bloc diagram of an example of process for preparing various products



FIG. 9 shows another bloc diagram of an example of process according to the present disclosure;



FIGS. 10A and 10B show further bloc diagrams of examples of processes according to the present disclosure;



FIGS. 11A and 11B show a further bloc diagrams of examples of processes according to the present disclosure; FIGS. 12A and 12B show further bloc diagrams of examples of processes according to the present disclosure;



FIG. 13 shows another bloc diagram of an example of process for preparing alumina and various other products according to the present disclosure;



FIG. 14 shows another bloc diagram of an example of process for preparing alumina and various other products according to the present disclosure;



FIG. 15 shows solubilisation curves of various metal chlorides as a function of HCl concentration;



FIG. 16 shows solubilisation curves of MgCl2 at various temperatures;



FIG. 17 shows solubilisation curves of various metal chlorides as a function of HCl concentration; and



FIG. 18 shows a bloc diagram of an example of process for preparing alumina and various other products according to the present disclosure.





DETAILED DESCRIPTION OF VARIOUS EMBODIMENTS

The following non-limiting examples further illustrate the technology described in the present disclosure.


The aluminum-containing material can be for example chosen from aluminum-containing ores (such as aluminosillicate minerals, clays, argillite, nepheline, mudstone, beryl, cryolite, garnet, spinel, bauxite, carbonatite, kyanite, kaolin, serpentine or mixtures thereof can be used). The aluminum-containing material can also be a recycled industrial aluminum-containing material such as slag, red mud or fly ashes.


The expression “red mud” as used herein refers, for example, to an industrial waste product generated during the production of alumina. For example, such a waste product can comprise silica, aluminum, iron, calcium, and optionally titanium. It can also comprise an array of minor constituents such as Na, K, Cr, V, Ni, Ba, Cu, Mn, Pb, and/or Zn etc. For example, red mud can comprises about 15 to about 80% by weight of Fe2O3, about 1 to about 35% by weight Al2O3, about 1 to about 65% by weight of SiO2, about 1 to about 20% by weight of Na2O, about 1 to about 20% by weight of CaO, and from 0 to about 35% by weight of TiO2. According to another example, red mud can comprise about 30 to about 65% by weight of Fe2O3, about 10 to about 20% by weight Al2O3, about 3 to about 50% by weight of SiO2, about 2 to about 10% by weight of Na2O, about 2 to about 8% by weight of CaO, and from 0 to about 25% by weight of TiO2.


The expression “fly ashes” as used herein refers, for example, to an industrial waste product generated in combustion. For example, such a waste product can contain various elements such as silica, oxygen, aluminum, iron, calcium. For example, fly ashes can comprise silicon dioxide (SiO2) and aluminium oxide (Al2O3). For example, fly ashes can further comprises calcium oxide (CaO) and/or iron oxide (Fe2O3). For example fly ashes can comprise fine particles that rise with flue gases. For example, fly ashes can be produced during combustion of coal. For example, fly ashes can also comprise at least one element chosen from arsenic, beryllium, boron, cadmium, chromium, chromium VI, cobalt, lead, manganese, mercury, molybdenum, selenium, strontium, thallium, and/or vanadium. For example, fly ashes can also comprise rare earth elements and rare metals. For example, fly ashes can be considered as an aluminum-containing material.


The expression “slag” as used herein refers, for example, to an industrial waste product comprising aluminum oxide and optionally other oxides such as oxides of calcium, magnesium, iron, and/or silicon.


The expression “rare earth element” (also described as “REE”) as used herein refers, for example, to a rare element chosen from scandium, yttrium, lanthanum, cerium, praseodymium, neodymium, promethium, samarium, europium, gadolinium, terbium, dysprosium, holmium, erbium, thulium, ytterbium, and lutetium. The expression “rare metals” as used herein refers, for example, to rare metals chosen from indium, zirconium, lithium, and gallium. These rare earth elements and rare metals can be in various form such as the elemental form (or metallic form), under the form of chlorides, oxides, hydroxides etc. The expression “rare earths” as used in the present disclosure as a synonym of “rare earth elements and rare metals” that is described above.


The expression “at least one iron chloride” as used herein refers to FeCl2, FeCl3 or a mixture thereof.


The term “hematite” as used herein refers, for example, to a compound comprising α-Fe2O3, γ-Fe2O3, β-FeO.OH or mixtures thereof.


The term “serpentine” as used herein refers, for example, to an ore that comprises Mg and optionally iron. For example, the ore can also comprise nickel, aluminum and/or cobalt. For example, the serpentine can be chosen from antigorite, chrysotile and lizardite.


The expression “iron ions” as used herein refers, for example to ions comprising to at least one type of iron ion chosen from all possible forms of Fe ions. For example, the at least one type of iron ion can be Fe2+, Fe3+, or a mixture thereof.


The expression “aluminum ions” as used herein refers, for example to ions comprising to at least one type of aluminum ion chosen from all possible forms of Al ions. For example, the at least one type of aluminum ion can be Al3+.


The expression “at least one aluminum ion”, as used herein refers, for example, to at least one type of aluminum ion chosen from all possible forms of Al ions. For example, the at least one aluminum ion can be Al3+.


The expression “at least one iron ion”, as used herein refers, for example, to at least one type of iron ion chosen from all possible forms of Fe ions. For example, the at least one iron ion can be Fe2+, Fe3+, or a mixture thereof.


The expression “at least one precipitated iron ion”, as used herein refers, for example, to at least one type of iron ion chosen from all possible forms of Fe ions that was precipitated in a solid form. For example, the at least one iron ion present in such a precipitate can be Fe2+, Fe3+, or a mixture thereof.


Terms of degree such as “about” and “approximately” as used herein mean a reasonable amount of deviation of the modified term such that the end result is not significantly changed. These terms of degree should be construed as including a deviation of at least ±5% or at least ±10% of the modified term if this deviation would not negate the meaning of the word it modifies.


The expression “substantially selectively isolate” as used herein when referring to isolating a compound refers, for example, to isolating such a compound together with less than 30, 25, 20, 15, 10, 5, 3, 2 or 1% of impurities. Such impurities can be other compounds such as other metals.


The expressions “substantially selectively precipitating”, “substantially selectively precipitate” and their equivalents as used herein when referring to precipitating a compound refers, for example, to precipitating such a compound together with less than 30, 25, 20, 15, 10, 5, 3, 2 or 1% of impurities. Such impurities can be other compounds such as other metals.


For example, the material can be leached with HCl having a concentration of about 10 to about 50 weight %, about 15 to about 45 weight %, of about 18 to about 45 weight % of about 18 to about 32 weight %, of about 20 to about 45 weight %, of about 25 to about 45 weight %, of about 26 to about 42 weight %, of about 28 to about 40 weight %, of about 30 to about 38 weight %, or between 25 and 36 weight %. For example, HCl at about 18 wt % or about 32 wt % can be used.


Leaching can also be carried out by adding dry highly concentrated acid (for example, 85%, 90% or 95%) in gas phase into the aqueous solution. Alternatively, leaching can also be carried out by using a weak acid solution (for example <3 wt %).


For example, leaching can be carried out by using HCl having a concentration of about 18 to about 32 wt % in a first reactor and then, by using HCl having concentration of about 90 to about 95%, or about 95 to about 100% (gaseous) in a second reactor.


For example, leaching can be carried out by using HCl having a concentration of about 18 to about 32 wt % in a first reactor then, by using HCl having concentration of about 90 to about 95% (gaseous) in a second reactor; and by using HCl having concentration of about 90 to about 95% (gaseous) in a third reactor.


For example, leaching can be carried out under an inert gas atmosphere (for example argon or nitrogen).


For example, leaching can be carried out under an atmosphere of NH3.


For example, the material can be leached at a temperature of about 125 to about 225° C., about 150 to about 200° C., about 160 to about 190° C., about 185 to about 190° C., about 160 to about 180° C., about 160 to about 175° C., or about 165 to about 170° C.


For example, the material can be leached at a pressure of about 4 to about 10 barg, about 4 to about 8 barg, or about 5 to about 6 barg.


For example a first leaching can be carried out at atmospheric pressure and then, at least one further leaching (for example 1, 2 or 3 subsequent leaching steps) can be carried out under pressure.


For example, leaching can be a continuous leaching or semi-continuous.


For example, the material can be an aluminum-containing material.


For example, the material can be an iron-containing material.


For example, the material can be a zinc-containing material.


For example, the material can be a copper-containing material.


For example, the material can be a gold-containing material.


For example, the material can be a silver-containing material.


For example, the material can be a molybdenum-containing material.


For example, the material can be a cobalt-containing material.


For example, the material can be a magnesium-containing material.


For example, the material can be a lithium-containing material.


For example, the material can be a manganese-containing material.


For example, the material can be a nickel-containing material.


For example, the material can be a palladium-containing material.


For example, the material can be a platinum-containing material.


For example, the material can be a magnesium-containing material.


For example, the material can be a lithium-containing material.


For example, the material can be a thorium-containing material.


For example, the material can be a phosphorus-containing material.


For example, the material can be a an uranium-containing material.


For example, the material can be a titanium-containing material.


For example, the material can be a rare earth elements-containing material.


For example, the material can be a rare metal-containing material.


The processes of the present disclosure can be effective for treating various materials. The at least one material can be an aluminum-containing material, The aluminum-containing material can be an aluminum-containing ore. For example, clays, argillite, mudstone, beryl, cryolite, garnet, spinel, bauxite, serpentine or mixtures thereof can be used as starting material. The aluminum-containing material can also be a recycled industrial aluminum-containing material such as slag. The aluminum-containing material can also be red mud.


The at least one material can be a nickel-containing material. The nickel-containing material can be a nickel-containing ore.


The at least one material can be a zinc-containing material. The zinc-containing material can be a zinc-containing ore.


The at least one material can be a copper-containing material. The copper-containing material can be a copper-containing ore.


The at least one material can be a titanium-containing material. The titanium-containing material can be a titanium-containing ore.


The at least one material can be a magnesium-containing material. The magnesium-containing material can be a magnesium-containing ore.


The processes of the present disclosure can be effective for treating various nickel-containing ores. For example, niccolite, kamacite, taenite, limonite, garnierite, laterite, pentlandite, serpentine, or mixtures thereof can be used.


The processes of the present disclosure can be effective for treating various zinc-containing ores. For example, smithsonite, warikahnite, sphalerite, serpentine or mixtures thereof can be used.


The processes of the present disclosure can be effective for treating various copper-containing ores. For example, copper-containing oxide ores, can be used. For example, chalcopyrite, chalcocite, covellite, bornite, tetrahedrite, malachite, azurite, cuprite, chrysocolla, or mixtures thereof can also be used.


The processes of the present disclosure can be effective for treating various titanium-containing ores. For example, ecandrewsite, geikielite, pyrophanite, ilmenite, or mixtures thereof can be used.


The processes of the present disclosure can be effective for treating various magnesium-containing ores. For example, the magnesium-containing ore can be chosen from serpentine, asbestos, antigorite, chrysotile, lizardite, brucite, magnesite, dolomite, kieserite, bischofite, langbeinite, epsomite, kainite, carnallite, astrakanite, laterite, geikielite and polyhalite.


For example, in the processes, the leachate can be treated with HCl that is in gaseous form.


For example, the processes can comprise reacting the leachate with gaseous HCl so as to obtain the liquid and the precipitate comprising the first metal under the form of a chloride.


For example, the processes can comprise reacting the leachate with dry gaseous HCl so as to obtain the liquid and the precipitate comprising the first metal under the form of a chloride.


For example, precipitating AlCl3 can comprise crystallizing AlCl3.6H2O.


For example, the processes can comprise reacting the leachate with acid of at least 30% wt. that was recovered, regenerated and/or purified as indicated in the present disclosure so as to obtain the liquid and the precipitate comprising the aluminum ions in the form of AlCl3.6H2O.


For example, the processes can further comprise recycling the gaseous HCl so-produced by contacting it with water so as to obtain a composition having a concentration of about 18 to about 45 weight % or 25 to about 45 weight %.


For example, the processes can further comprise recycling the gaseous HCl so-produced by contacting it with water so as to obtain a composition having a concentration of about 18 to about 45 weight % or about 25 to about 45 weight % and using the composition for leaching the material.


For example, the liquid can comprise iron chloride. Iron chloride can comprise at least one of FeCl2, FeCl3, and a mixture thereof.


For example, the liquid can have an iron chloride concentration of at least 30% by weight; and can then be hydrolyzed at a temperature of about 155 to about 350° C.


For example, the liquid can be concentrated to a concentrated liquid having an iron chloride concentration of at least 30% by weight; and then the iron chloride can be hydrolyzed at a temperature of about 155 to about 350° C. while maintaining a ferric chloride concentration at a level of at least 65% by weight, to generate a composition comprising a liquid and precipitated hematite, and recovering the hematite.


For example, non-hydrolysable elements with hematite can be concentrated back to a concentration of about 0.125 to about 52% wt. in circulation loop in view of selective extraction.


For example, the liquid can be concentrated to a concentrated liquid having a concentration of the at least one iron chloride of at least 30% by weight; and then hydrolyzed at a temperature of about 155 to about 350° C.


For example, the liquid can be concentrated to a concentrated liquid having a concentration of the at least one iron chloride of at least 30% by weight; and then the at least one iron chloride is hydrolyzed at a temperature of about 155 to about 350° C. while maintaining a ferric chloride concentration at a level of at least 65% by weight, to generate a composition comprising a liquid and precipitated hematite, and recovering the hematite.


For example, the liquid can be concentrated to a concentrated liquid having a concentration of the at least one iron chloride of at least 30% by weight; and then the at least one iron chloride is hydrolyzed at a temperature of about 155 to about 350° C. while maintaining a ferric chloride concentration at a level of at least 65% by weight, to generate a composition comprising a liquid and precipitated hematite; recovering the hematite; and recovering rare earth elements and/or rare metals from the liquid.


For example, the at least one iron chloride can be hydrolyzed at a temperature of about, 150 to about 175, 160 to about 175, 155 to about 170, 160 to about 170 or 165 to about 170° C.


For example, the liquid can be concentrated to a concentrated liquid having an iron chloride concentration of at least 30% by weight; and then the iron chloride can be hydrolyzed at a temperature of about 155 to about 350° C. while maintaining a ferric chloride concentration at a level of at least 65% by weight, to generate a composition comprising a liquid and precipitated hematite; recovering the hematite; and recovering rare earth elements and/or rare metals from the liquid.


For example, the processes can further comprise, after recovery of the rare earth elements and/or rare metals, reacting the liquid with HCl so as to cause precipitation of MgCl2, and recovering same.


For example, the processes can further comprise calcining MgCl2 into MgO.


For example, the processes can further comprises, after recovery of the rare earth elements and/or rare metals, reacting the liquid with HCl, and substantially selectively precipitating Na2SO4. For example, Na2SO4 can be precipitated by reacting the liquid with H2SO4.


For example, the processes can further comprises, after recovery of the rare earth elements and/or rare metals, reacting the liquid with HCl, and substantially selectively precipitating K2SO4. For example, K2SO4 can be precipitated by adding H2SO4.


For example, the liquid can be concentrated to a concentrated liquid having an iron chloride concentration of at least 30% by weight; and then the iron chloride can be hydrolyzed at a temperature of about 155 to about 350° C. while maintaining a ferric chloride concentration at a level of at least 65% by weight, to generate a composition comprising a liquid and precipitated hematite; recovering the hematite; and reacting the liquid with HCl. For example, such processes can further comprises reacting the liquid with H2SO4 so as to substantially selectively precipitate Na2SO4. The processes can also comprise further reacting the liquid with H2SO4 so as to substantially selectively precipitating K2SO4.


For example, the processes can comprise reacting dry individual salts (for example Na or K salts) obtained during the processes with H2SO4 and recovering HCl while producing marketable K2SO4 and Na2SO4 and recovering hydrochloric acid of about 15 to about 90% wt.


For example, sodium chloride produced in the processes can undergo a chemical reaction with sulfuric acid so as to obtain sodium sulfate and regenerate hydrochloric acid. Potassium chloride can undergo a chemical reaction with sulfuric acid so as to obtain potassium sulfate and regenerate hydrochloric acid. Sodium and potassium chloride brine solution can alternatively be the feed material to adapted small chlor-alkali electrolysis cells. In this latter case, common bases (NaOH and KOH) and bleach (NaOCl and KOCl) are produced.


For example, the processes can further comprise, after recovery of the rare earth elements and/or rare metals, recovering NaCl from the liquid, reacting the NaCl with H2SO4, and substantially selectively precipitating Na2SO4.


For example, the processes can further comprise, downstream of recovery of the rare earth elements and/or rare metals, recovering KCl from the liquid, reacting the KCl with H2SO4, and substantially selectively precipitating K2SO4.


For example, the processes can further comprise, downstream of recovery of the rare earth elements and/or rare metals, recovering NaCl from the liquid, carrying out an electrolysis to generate NaOH and NaOCl.


For example, the processes can further comprise, downstream of recovery of the rare earth elements and/or rare metals, recovering KCl from the liquid, reacting the KCl, carrying out an electrolysis to generate KOH and KOCl.


For example, the liquid can be concentrated to a concentrated liquid having a concentration of the at least one iron chloride of at least 30% by weight; and then the at least one iron chloride is hydrolyzed at a temperature of about 155 to about 350° C. while maintaining a ferric chloride concentration at a level of at least 65% by weight, to generate a composition comprising a liquid and precipitated hematite; recovering the hematite; and extracting NaCl and/or KCl from the liquid.


For example, the processes can further comprise reacting the NaCl with H2SO4 so as to substantially selectively precipitate Na2SO4.


For example, the processes can further comprise reacting the KCl with H2SO4 so as to substantially selectively precipitate K2SO4.


For example, the processes can further comprise carrying out an electrolysis of the NaCl to generate NaOH and NaOCl.


For example, the processes can further comprise carrying out an electrolysis of the KCl to generate KOH and KOCl.


For example, the processes can comprise separating the solid from the leachate and washing the solid so as to obtain silica having a purity of at least 95%, at least 96%, at least 97%, at least 98%, at least 99%, at least 99.5% or at least 99.9%.


For example, the processes can comprise reacting the leachate with gaseous HCl so as to obtain the liquid and the precipitate comprising the aluminum ions in the form of AlCl3.6H2O.


For example, the processes can comprise reacting the leachate with dry gaseous HCl so as to obtain the liquid and the precipitate comprising the aluminum ions in the form of AlCl3.6H2O.


For example, the processes can comprise reacting the leachate with acid of at least 30% wt. that was recovered, regenerated and/or purified as indicated in the present disclosure so as to obtain the liquid and the precipitate comprising the aluminum ions in the form of AlCl3.6H2O.


For example, the processes can comprise reacting the leachate with gaseous HCl so as to obtain the liquid and the precipitate comprising the aluminum ions, the precipitate being formed by crystallization of AlCl3.6H2O.


For example, the processes can comprise reacting the leachate with dry gaseous HCl so as to obtain the liquid and the precipitate comprising the aluminum ions, the precipitate being formed by crystallization of AlCl3.6H2O.


For example, aluminum ions can be precipitated under the form of AlCl3 (for example AlCl3.6H2O) in a crystallizer, for example, by adding HCl having a concentration of about 26 to about 32 wt %.


For example, the gaseous HCl can have a HCl concentration of at least 85% wt. or at least 90% wt.


For example, the gaseous HCl can have a HCl concentration of about 90% wt. about 90% to about 95% wt., or about 90% to about 99% wt.


For example, during the crystallization of AlCl3.6H2O, the liquid can be maintained at a concentration of HCl of about 25 to about 35% by weight or about 30 to about 32% by weight.


For example, the crystallization can be carried out at a temperature of about 45 to about 65° C. or about 50 to about 60° C.


For example, the HCl can be obtained from the gaseous HCl so-produced.


For example, in the processes of the present disclosure, a given batch or quantity of the material will be leached, will then be converted into AlCl3 and when the HCl generated during calcination of AlCl3 into Al2O3 will be used for example to leach another given batch or quantity of the material.


For example, the processes can comprise heating the precipitate at a temperature of at least 180, 230, 250, 300, 350, 400, 450, 500, 550, 600, 650, 700, 750, 800, 850, 900, 925, 930, 1000, 1100, 1200 or 1250° C. for converting AlCl3 or Al(OH)3 into Al2O3.


For example, converting AlCl3 into Al2O3 can comprise calcination of AlCl3.


For example, calcination is effective for converting AlCl3 into beta-Al2O3.


For example, calcination is effective for converting AlCl3 into alpha-Al2O3.


For example, converting AlCl3 into Al2O3 can comprise carrying out a calcination via a two-stage circulating fluid bed reactor.


For example, converting AlCl3 into Al2O3 can comprise carrying out a calcination via a two-stage circulating fluid bed reactor that comprises a preheating system.


For example, converting AlCl3 into Al2O3 can comprise carrying out a calcination at low temperature, for example, about 300 to about 600° C., about 325 to about 550° C., about 350 to about 500° C., about 375 to about 450° C., about 375 to about 425° C., or about 385 to about 400° C. and/or injecting steam.


For example, converting AlCl3 into Al2O3 can comprise carrying out a calcination at low temperature, for example, at least 180° C., at least 250° C., at least 300° C., at least 350° C. and/or injecting steam.


For example, converting AlCl3 into Al2O3 can comprise carrying out a calcination at low temperature, for example, less than 600° C. and/or injecting steam.


For example, converting AlCl3 into Al2O3 can comprise carrying out a calcination by using coal as combustion source and by using a degasification unit.


For example, steam (or water vapor) can be injected at a pressure of about 200 to about 700 psig, about 300 to about 700 psig, about 400 to about 700 psig, about 550 to about 650 psig, about 575 to about 625 psig, or about 590 to about 610 psig.


For example, steam (or water vapor) can be injected and a plasma torch can be used for carrying fluidization.


For example, the steam (or water vapor) can be overheated.


For example, the steam (or water vapor) can be at a temperature of about 300 to about 400° C.


For example, acid from the offgases generated during calcination can be then treated via a gas phase purification process.


For example, converting AlCl3 into Al2O3 can comprise carrying out a calcination by means of carbon monoxide (CO).


For example, converting AlCl3 into Al2O3 can comprise carrying out a calcination by means of a Refinery Fuel Gas (RFG).


For example, calcination can be carried out by injecting water vapor or steam and/or by using a combustion source chosen from fossil fuels, carbon monoxide, a Refinery Fuel Gas, coal, or chlorinated gases and/or solvants.


For example, calcination can be carried out by injecting water vapor or steam and/or by using a combustion source chosen from natural gas or propane.


For example, calcination can be carried out by providing heat by means of electric heating, gas heating, microwave heating.


The obtained alumina can be washed by demineralized water so as to at least partially remove NaCl and/or KCl.


For example, the fluid bed reactor can comprise a metal catalyst chosen from metal chlorides.


For example, the fluid bed reactor can comprise a metal catalyst that is FeCl3, FeCl2 or a mixture thereof.


For example, the fluid bed reactor can comprise a metal catalyst that is FeCl3.


For example, the preheating system can comprise a plasma torch.


For example, steam can be used as the fluidization medium heating. Heating can also be electrical.


For example, a plasma torch can be used for preheating the calcination reactor.


For example, a plasma torch can be used for preheating air entering in the calcination reactor.


For example, a plasma torch can be used for preheating a fluid bed.


For example, the calcination medium can be substantially neutral in terms of O2 (or oxidation). For example, the calcination medium can favorize reduction (for example a concentration of CO of about 100 ppm).


For example, the calcination medium is effective for preventing formation of Cl2.


For example, the processes can comprise converting AlCl3.6H2O into Al2O3 by carrying out a calcination of AlCl3.6H2O that is provided by the combustion of gas mixture that comprises:

    • CH4: 0 to about 1% vol;
    • C2H6: 0 to about 2% vol;
    • C3H8: 0 to about 2% vol;
    • C4H10: 0 to about 1% vol;
    • N2: 0 to about 0.5% vol;
    • H2: about 0.25 to about 15.1% vol;
    • CO: about 70 to about 82.5% vol; and
    • CO2: about 1.0 to about 3.5% vol.


Such a mixture can be efficient for reduction in off gas volume of 15.3 to 16.3%; therefore the capacity increases of 15.3 to 16.3% proven on practical operation of the circulating fluid bed. Thus for a same flow it represents an Opex of 0.65*16.3%=10.6%.


For example, the air to natural gas ratio of (Nm3/h over Nm3/h) in the fluid bed can be about 9.5 to about 10


For example, the air to CO gas ratio of (Nm3/h over Nm3/h) in the fluid bed can be about 2 to about 3.


For example, the processes can comprise, before leaching the material, a pre-leaching removal of fluorine optionally contained in the material.


For example, the processes can comprise leaching of the material with HCl so as to obtain the leachate comprising aluminum ions and the solid, separating the solid from the leachate; and further treating the solid so as to separate SiO2 from TiO2 that are contained therein.


For example, the processes can comprise leaching the material with HCl so as to obtain the leachate comprising aluminum ions and the solid, separating the solid from the leachate; and further treating the solid with HCl so as to separate Si from Ti that are contained therein.


For example, the processes can comprise leaching the material with HCl so as to obtain the leachate comprising aluminum ions and the solid, separating the solid from the leachate; and further treating the solid with HCl at a concentration of less than 20% wt., at a temperature of less than 85° C., in the presence of MgCl2, so as to separate Si from Ti that are contained therein.


For example, converting AlCl3 into Al2O3 can comprise carrying out a one-step calcination.


For example, calcination can be carried out at different temperatures with steam. Temperature applied of superheated steam can be of about 350° C. to about 550° C. or about 350° C. to about 940° C. or about 350° C. to about 1200° C.


For example, multi stage evaporation step of the hydrolyser can be carried out to reduce drastically energy consumption.


For example, the processes can be effective for providing an Al2O3 recovery yield of at least 93%, at least 94%, at least 95%, about 90 to about 95%, about 92 to about 95%, or about 93 to about 95%.


For example, the processes can be effective for providing a Fe2O3 recovery yield of at least 98%, at least 99%, about 98 to about 99.5%, or about 98.5 to about 99.5%.


For example, the processes can be effective for providing a MgO recovery yield of at least 96%, at least 97%, at least 98%, or about 96 to about 98%.


For example, the processes can be effective for providing a HCl recovery yield of at least 98%, at least 99%, or about 98 to about 99.9%.


For example, the processes can be effective for providing chlorides of rare earth elements (REE-Cl) and chlorides of rare metals (RM-Cl) in recovery yields of about 75% to about 96.5% by using internal processes via an internal concentration loop.


For example, the processes can be effective for providing hydrochloric acid recovery yield of about 99.75% with non-hydrolysable elements.


For example, the material can be argillite.


For example, the material can be bauxite.


For example, the material can be red mud.


For example, the material can be fly ashes.


For example, the material can be chosen from industrial refractory materials.


For example, the material chosen from aluminosilicate minerals.


For example, the processes can be effective for avoiding producing red mud.


For example, the alumina and the other products are substantially free of red mud.


For example, HCl can be recycled. For example, such a recycled HCl can be concentrated and/or purified.


For example, gaseous HCl can be concentrated and/or purified by means of H2SO4. For example, gaseous HCl can be passed through a packed column where it is contacted with a H2SO4 countercurrent flow. For example, by doing so, concentration of HCl can be increased by at least 50% wt., at least 60% wt., at least 70% wt., at least 75% wt., at least 80% wt., about 50% wt. to about 80% wt., about 55% wt. to about 75% wt., or about 60% wt. For example, the column can be packed with a polymer such as polypropylene(PP) or polytrimethylene terephthalate (PTT).


For example, gaseous HCl can be concentrated and/or purified by means of CaCl2 or LiCl. For example, gaseous HCl can be passed through a column packed with CaCl2 or LiCl.


For example, AlCl3.6H2O obtained in the processes of the present disclosure can be further purified as described in U.S. 61/726,079, that is hereby incorporated by reference in its entirety.


For example, MgCl2 can be substantially selectively precipitated from the leachate and removed therefrom and then, the leachate can be reacted with HCl so as to obtain the liquid and the precipitate comprising the aluminum ions in the form of AlCl3, and separating the precipitate from the liquid.


For example, the leachate can be reacted with HCl so as to obtain the liquid and the precipitate comprising the aluminum ions in the form of AlCl3, and separating the precipitate from the liquid, and then the MgCl2 is substantially selectively precipitated from the leachate and removed therefrom.


For example, the aluminum-containing material can bleached with HCl so as to obtain the leachate comprising aluminum ions, magnesium ions and the solid, and the solid is separated from the leachate at a temperature of at least 50, 60, 75 or 100° C. For example, a filtration can be carried out and the temperature of the leachate can have a value as previously indicated.


For example, MgCl2 can be substantially selectively precipitated from the leachate at a temperature of about 5 to about 70° C., about 10 to about 60° C., about 10 to about 40° C., or about 15 to about 30° C.


For example, the processes can comprise, before reacting the leachate with HCl so as to obtain the liquid and the precipitate, controlling the temperature of the leachate so as to substantially selectively precipitate a second metal in the form of a chloride, and removing the precipitate from the leachate.


For example, the processes can comprise, after reacting the leachate with HCl so as to obtain the liquid and the precipitate, controlling the temperature of the leachate so as to substantially selectively precipitate a second metal in the form of a chloride, and removing the precipitate from the leachate.


For example, the processes can further comprises treating the precipitate under conditions effective for converting the chloride of the first metal it into an oxide of the first metal and optionally recovering gaseous HCl so-produced.


For example, the processes can further comprises treating the precipitate under conditions effective for converting the chloride of the second metal it into an oxide of the second metal and optionally recovering gaseous HCl so-produced.


For example, the solid can be treated with HCl and the metal chloride so as to obtain a liquid portion comprising Ti and a solid portion containing Si and wherein the liquid portion is separated from the solid portion.


For example, the solid can be treated with HCl and the metal chloride so as to obtain a liquid portion comprising TiCl4.


For example, the process can further comprise converting TiCl4 into TiO2.


For example, TiCl4 can be converted into TiO2 by solvent extraction of the third liquid fraction and subsequent formation of titanium dioxide from the solvent extraction.


For example, TiCl4 can be reacted with water and/or a base to cause precipitation of TiO2.


For example, TiCl4 can be converted into TiO2 by means of a pyrohydrolysis, thereby generating HCl.


For example, TiCl4 can be converted into TiO2 by means of a pyrohydrolysis, thereby generating HCl that is recycled.


For example, the metal chloride can be MgCl2 or ZnCl2.


For example, the solid can comprise TiO2 and SiO2 and the solid is treated with Cl2 and carbon in order to obtain a liquid portion and a solid portion, and wherein the solid portion and the liquid portion are separated from one another.


For example, the liquid portion can comprise TiCl2 and/or TiCl4.


For example, the liquid portion can comprise TiCl4.


For example, the process can further comprise heating TiCl4 so as to convert it into TiO2.


For example, the obtained TiO2 can be purified by means of a plasma torch.


For example, the various products obtained by the processes of the present disclosure such as alumina, hematite, titanium oxides, magnesium oxides, rare earth elements and rare metals can be further purified by means of a plasma torch. For example, the rare earth elements and rare metals, once isolated, can be individually injected into a plasma torch so as to further purify them.


For example, the processes can further comprise converting alumina (Al2O3) into aluminum. Conversion of alumina into aluminum can be carried out, for example, by using the Hall-Héroult process. References is made to such a well known process in various patents and patent applications such as US 20100065435; US 20020056650; U.S. Pat. No. 5,876,584; U.S. Pat. No. 6,565,733. Conversion can also be carried out by means of other methods such as those described in U.S. Pat. No. 7,867,373; U.S. Pat. No. 4,265,716; U.S. Pat. No. 6,565,733 (converting alumina into aluminum sulfide followed by the conversion of aluminum sulfide into aluminum.). For example, aluminium can be produced by using a reduction environment and carbon at temperature below 200° C. Aluminum can also be produced by reduction using potassium and anhydrous aluminum chloride (Wohler Process).


For example, controlling the temperature of the leachate so as to precipitate the the first metal in the form of a chloride, and removing the precipitate from the leachate, can be carried out before reacting the leachate with HCl so as to obtain a liquid and a precipitate comprising a chloride of the second metal, and separating the precipitate from the liquid.


For example, controlling the temperature of the leachate so as to precipitate the the first metal in the form of a chloride, and removing the precipitate from the leachate, can be carried out after reacting the leachate with HCl so as to obtain a liquid and a precipitate comprising a chloride of the second metal, and separating the precipitate from the liquid.


For example, reacting the leachate with HCl so as to obtain a precipitate comprising the first metal in the form of a chloride, can be carried out by substantially selectively precipitating the first metal chloride.


For example, reacting the leachate with HCl so as to obtain a precipitate comprising the second metal in the form of a chloride, can be carried out by substantially selectively precipitating the second metal chloride,


For example, controlling the temperature of the leachate so as to precipitate the the first metal in the form of a chloride can be carried out substantially selectively.


For example, controlling the temperature of the leachate so as to precipitate the second metal in the form of a chloride can be carried out substantially selectively.


For example, controlling the concentration of HCl in the leachate and/or the temperature of the leachate so as to precipitate the first metal in the form of a chloride, can be carried out substantially selectively.


For example, controlling the concentration of HCl in the leachate and/or the temperature of the leachate so as to precipitate the second metal in the form of a chloride, can be carried out substantially selectively.


For example, the first metal can chosen from aluminum, iron, zinc, copper, gold, silver, molybdenum, cobalt, magnesium, lithium, manganese, nickel, palladium, platinum, thorium, phosphorus, uranium and titanium, and/or at least one rare earth element and/or at least one rare metal


For example, the liquid can comprise a second metal.


For example, the second metal can be chosen from aluminum, iron, zinc, copper, gold, silver, molybdenum, cobalt, magnesium, lithium, manganese, nickel, palladium, platinum, thorium, phosphorus, uranium and titanium, and/or at least one rare earth element and/or at least one rare metal


For example, the process can comprise separating the precipitate from the liquid and heating the second metal in order to convert a chloride of the second metal into an oxide of the second metal.


For example, the second metal can be magnesium.


For example, the second metal can be aluminum.


For example, the first metal can be aluminum and the second metal can be magnesium.


For example, the second metal can be aluminum and the first metal can be magnesium.


For example, the processes can comprise:

    • separating the solid from the leachate;
    • leaching the solid with an acid so as to obtain another leachate; and
    • recovering a third metal from the another leachate.


For example, the third metal can be chosen from aluminum, iron, zinc, copper, gold, silver, molybdenum, cobalt, magnesium, lithium, manganese, nickel, palladium, platinum, thorium, phosphorus, uranium and titanium, and/or at least one rare earth element and/or at least one rare metal.


For example, the third metal can be titanium.


For example, the acid can be chosen from HCl, HNO3, H2SO4 and mixtures thereof.


For example, the process can comprise recovering the third metal from the another leachate by precipitating the third metal.


For example, the third metal can be precipitated by reacting it with HCl.


For example, the process can further comprise heating the third metal in order to convert a chloride of the third metal into an oxide of the third metal.


For example, the first metal can be magnesium.


For example, the first metal can be nickel.


For example, the second metal can be magnesium.


For example, the second metal can be nickel.


For example, the process can comprise reacting the leachate with gaseous HCl so as to obtain a liquid and a precipitate comprising MgCl2.


For example, the process comprises reacting the leachate with gaseous HCl so as to obtain a liquid and a precipitate comprising MgCl2.


For example, NaCl recovered from the processes of the present disclosure can be reacted with SO2, so as to produce HCl and Na2SO4. Such a reaction that is an exothermic reaction can generate steam that can be used to activate a turbine and eventually produce electricity.


For example, U and/or Th can be treated with the processes of the present disclosure. For example, these two elements can be in such processes in admixtures with iron ions and they can be separated therefrom by means of at least one ion exchange resin.


For example, the processes can comprise substantially selectively precipitating the magnesium ions by reacting the leachate with the precipitating agent.


For example, the precipitating agent can be Mg(OH)2.


For example, the at least one metal can be nickel.


For example, the at least one metal can be cobalt.


For example, the at least one metal can be iron.


For example, the at least one metal can be aluminum.


In the processes of the present disclosure, when the material to be treated comprises aluminum and magnesium, magnesium can be first removed from the leachate by controlling temperature of said leachate so as to substantially selectively cause precipitation (or crystallization) of MgCl2, remove it from the leachate and then substantially selectively cause precipitation of AlCl3 by reacting the leachate with HCl (for example gaseous HCl). Alternatively, the leachate can be reacted with HCl to substantially selectively cause precipitation (or crystallization) of AlCl3 (for example gaseous HCl). In such a case the temperature can be maintained for example above 50, 60, 70, 80, or 90° C. AlCl3 is then removed from the leachate and then, temperature of the leachate is controlled so as to substantially selectively cause precipitation of MgCl2. Depending on the concentration of Al vs Mg in the starting material one scenario or the other can be selected. For example, if the concentration of Mg is greater than the concentration of Al, Mg can be removed first from the leachate. For example, if the concentration of Al is greater than the concentration of Mg, Al can be removed first from the leachate.


In the processes of the present disclosure, when the material to be treated comprises aluminum, iron and magnesium. Magnesium can be first removed from the leachate by controlling temperature of said leachate so as to substantially selectively cause precipitation (or crystallization) of MgCl2, remove it from the leachate and then substantially selectively cause precipitation of AlCl3 by reacting the leachate with HCl (for example gaseous HCl). Then, the remaining composition comprising iron chloride can be treated so as to convert iron chloride into iron oxide by using one of the methods discussed in the present disclosure. Alternatively, the leachate can be reacted with HCl to substantially selectively cause precipitation (or crystallization) of AlCl3 (for example gaseous HCl). In such a case the temperature can be maintained for example above 50, 60, 70, 80, or 90° C. AlCl3 is then removed from the leachate and then, temperature of the leachate is controlled so as to substantially selectively cause precipitation of MgCl2. Then, the remaining composition comprising iron chloride can be treated so as to convert iron chloride into iron oxide by using the methods discussed in the present disclosure.


For example, the precipitate can be reacted with a base (for example KOH or NaOH). For example, AlCl3 can be converted into Al(OH)3 before calcination.


According to one example as shown in FIG. 1, the processes can involve the following steps (the reference numbers in FIG. 1 correspond to the following steps):



1—The aluminum-containing material is reduced to an average particle size of about 50 to about 80 μm.



2—The reduced and classified material is treated with hydrochloric acid which allows for dissolving, under a predetermined temperature and pressure, the aluminum with other elements like iron, magnesium and other metals including rare earth elements and/or rare metals. The silica and titanium (if present in raw material) remain totally undissolved.



3—The mother liquor from the leaching step then undergoes a separation, a cleaning stage in order to separate the solid from the metal chloride in solution.



4—The spent acid (leachate) obtained from step 3 is then brought up in concentration with dry and highly concentrated gaseous hydrogen chloride by sparging this one into a crystallizer. This results into the crystallization of aluminum chloride hexahydrate (precipitate) with a minimum of other impurities. Depending on the concentration of iron chloride at this stage, further crystallization step(s) can be required. The precipitate is then separated from the liquid. For example, particle size of crystals can be about 100 to about 500 microns, about 200 to about 400 microns, or about 200 to about 300 microns. Alternatively, particle size of crystals can be about 100 to about 200 microns, about 300 to about 400 microns or about 400 to 500 microns.



5—The aluminum chloride hexahydrate is then calcined (for example by means of a rotary kiln, fluid bed, etc) at high temperature in order to obtain the alumina form. Highly concentrated gaseous hydrogen chloride is then recovered and excess is brought in aqueous form to the highest concentration possible so as to be used (recycled) in the acid leaching step. Acid can also be directly sent in gas phase to the acid purification stage to increase HCl concentration from about 30 wt % to about 95 wt %. This can be done, for example, during drying stage.



6—Iron chloride (the liquid obtained from step 4) is then pre-concentrated and hydrolyzed at low temperature in view of the Fe2O3 (hematite form) extraction and acid recovery from its hydrolysis. All heat recovery from the calcination step (step 5), the leaching part exothermic reaction (step 1) and other section of the processes is being recovered into the pre-concentrator.



10—After the removal of hematite, a solution rich in rare earth elements and/or rare metals can be processed. As it can be seen in FIG. 3, an internal recirculation can be done (after the removal of hematite) and the solution rich in rare earth elements and/or rare metals can be used for crystallization stage 4. Extraction of the rare earth elements and/or rare metals can be done as described in WO/2012/126092 and/or WO/2012/149642. These two documents are hereby integrated by reference in their entirety.


Other non-hydrolysable metal chlorides (Me—Cl) such as MgCl2 and others then undergo the following steps:



7—The solution rich in magnesium chloride and other non-hydrolysable products at low temperature is then brought up in concentration with dry and highly concentrated gaseous hydrogen chloride by sparging it into a crystallizer. This results into the precipitation of magnesium chloride as an hexahydrate, for example after sodium and potassium chloride removal.



8—Magnesium chloride hexahydrate is then calcined (either through a rotary kiln, fluid bed, etc.) and hydrochloric acid at very high concentration is thus regenerated and brought back to the leaching step.



9—Other Me—Cl undergo a standard pyrohydrolysis step where mixed oxides (Me-O) can be produced and hydrochloric acid at the azeotropic point (20.2% wt.) is regenerated.



11—Ti contained in the solid obtained from step 3 can be treated so as to separate Si from Ti and thus obtain SiO2 and TiO2.


NaCl produced in this process can undergo chemical reaction with H2SO4 to produce Na2SO4 and HCl at a concentration at or above azeotropic concentration. Moreover, KCl can undergo chemical reaction with H2SO4 to produce K2SO4 and HCl having a concentration that is above the azeotropic concentration. Sodium and potassium chloride brine solution can be the feed material to adapted small chlor-alkali electrolysis cells. In this latter case, common bases (NaOH and KOH) and bleach (NaOCl and KOCl) are produced as well as HCl.


For example, the liquid can be concentrated to a concentrated liquid having an iron chloride concentration of at least 30% by weight; and then the iron chloride can be hydrolyzed at a temperature of about 155 to about 350° C. while maintaining a ferric chloride concentration at a level of at least 65% by weight, to generate a composition comprising a liquid and precipitated hematite, and recovering the hematite.


For example, the liquid can be concentrated to a concentrated liquid having an iron chloride concentration of at least 30% by weight; and then the iron chloride can be hydrolyzed at a temperature of about 155 to about 350° C. while maintaining a ferric chloride concentration at a level of at least 65% by weight, to generate a composition comprising a liquid and precipitated hematite; recovering the hematite; and recovering rare earth elements and/or rare metals from the liquid. For example, the process can further comprise, after recovery of the rare earth elements and/or rare metals, reacting the liquid with HCl so as to cause precipitation of MgCl2, and recovering same.


As previously indicated, various aluminum-containing materials can be used as starting material of the processes disclosed in the present disclosure. Examples with clays and bauxite have been carried out. However, the person skilled in the art will understand that the continuous processes can handle high percentages of silica (>55%) and impurities as well as relatively low percentages of aluminum (for example as low as about 15%) and still being economically and technically viable. Satisfactory yields can be obtained (>93-95%) on Al2O3 and greater than 75%, 85 or 90% on rare earth elements and/or rare metals. No pre-thermal treatment in most cases are required. The processes disclosed in the present disclosure involve special techniques on leaching and acid recovery at very high strength, thereby offering several advantages over alkaline processes.


In step 1 the mineral, whether or not thermally treated is crushed, milled, dried and classified to have an average particle size of about 50 to about 80 μm.


In step 2, the milled raw material is introduced into the reactor and will undergo the leaching phase.


The leaching hydrochloric acid used in step 2 can be a recycled or regenerated acid from steps 5, 6, 8, 9, 10 and 11 (see FIG. 3) its concentration can vary from 15% to 45% weight. percent. Higher concentration can be obtained using membrane separation, cryogenic and/or high pressure approach. The acid leaching can be carried out under pressure and at temperature close to its boiling point thus, allowing a minimal digestion time and extended reaction extent (90%-100%). Leaching (step 2) can be accomplished in a semi-continuous mode where spent acid with residual free hydrochloric acid is replaced by highly concentrated acid at a certain stage of the reaction or allowing a reduced acid/mineral ratio, thereby reducing reaction time and improving reaction kinetics. For example, kinetic constant k can be: 0.5-0.75 g/mole.L. For example, leaching can be continuous leaching.


As previously indicated, alkali metals, iron, magnesium, sodium, calcium, potassium, rare earth elements and other elements will also be in a chloride form at different stages. Silica and optionally titanium can remain undissolved and will undergo (step 3) a liquid/solid separation and cleaning stage. The processes of the present disclosure tend to recover maximum amount of free hydrochloric acid left and chlorides in solution in order to maximize hydrochloric acid recovery yield, using techniques such as rake classifying, filtration with band filters, centrifugation, high pressure, rotofilters and others. Thanks to step 13, Ti contained in the solid obtained from step 3 can be treated so as to separate Si from Ti and thus obtain SiO2 and TiO2. Various possible strategies can be used to separated Si from Ti as previously indicated. For example, the solid can be further leached (for example with HCl in the presence of a metal chloride (for example MgCl2 or ZnCl2) so as to solubilize Ti (for example in the form of TiCl4) while the Si remains solid. Alternatively, the solid can be reacted with Cl2 (see FIGS. 10A and 10B). The purified silica can then optionally undergo one or two additional leaching stages (for example at a temperature of about 150 to about 160° C.) so as to increase the purity of silica above 99.9%.


Pure SiO2 (one additional leaching stage) cleaning with nano water purity 99% min. Mother liquor free of silica is then named as spent acid (various metal chlorides and water) and goes to the crystallization step (step 4). Free HCl and chlorides recovery can be at least 99, 99.5 or 99.9%


In step 4, the spent acid (or leachate) with a substantial amount of aluminum chloride is then saturated with dry and highly concentrated gaseous hydrogen chloride obtained or recycled from step 5 or with aqueous HCl >30% wt., which results in the precipitate of aluminum chloride hexahydrate (AlCl3.6H2O). The precipitate retained is then washed and filtered or centrifuged before being fed to the calcination stage (step 5). The remaining of the spent acid from step 4 is then processed to acid recovery system (steps 6 to 8) where pure secondary products will be obtained.


In step 5, aluminum oxide (alumina) is directly obtained from high temperature conditions. The highly concentrated hydrogen chloride in gaseous form obtained can be fed to steps 4 and 7 for crystallization where it can be treated through hydrophobic membranes. The excess hydrogen chloride is absorbed and used as regenerated acid to the leaching step 2 as highly concentrated acid, higher than the concentration at the azeotropic point (>20.2%). For example, such a concentration can be about 18 to about 45 weight %, about 25 to about 45 weight % or between 25 and 36 weight %. Acid can also be redirected in gas phase directly (>30 wt %) to acid purification.


After step 4, various chlorides derivatives (mainly iron with magnesium and rare earth elements and rare metals) are next subjected to an iron extraction step. Such a step can be carried out for example by using the technology disclosed in WO 2009/153321, which is hereby incorporated by reference in its entirety. Moreover, hematite can be seeded for crystal growth. For example, hematite seeding can comprise recirculating the seeding.


In step 6, a hydrolysis at low temperature (155-350° C.) is carried out and pure Fe2O3 (hematite) is being produced and hydrochloric acid of at least 15% concentration is being regenerated. The method as described in WO 2009/153321 is processing the solution of ferrous chloride and ferric chloride, possible mixtures thereof, and free hydrochloric acid through a series of steps pre-concentration step, oxidation step where ferrous chloride is oxidized into ferric form, and finally through an hydrolysis step into an operational unit called hydrolyser where the ferric chloride concentration is maintained at 65 weight % to generate a rich gas stream where concentration ensures a hydrogen chloride concentration of 15-20.2% and a pure hematite that will undergo a physical separation step. Latent heat of condensation is recovered to the pre-concentration and used as the heating input with excess heat from the calcination stage (step 5).


The mother liquor from the hydrolyser (step 6) can be recirculated partially to first step crystallization process where an increase in concentration of non-hydrolysable elements is observed. After iron removal, the liquor is rich in other non-hydrolysable elements and mainly comprises magnesium chloride or possible mixture of other elements (various chlorides) and rare earth elements and rare metals that are, for example, still in the form of chlorides.


Rare earth elements and rare metals in form of chlorides are highly concentrated, in percentage, into the hydrolyser operational unit (step 6) and are extracted from the mother liquor (step 10) where various known techniques can be employed to extract a series of individual RE-O (rare earth oxides). Among others, the processes of the present disclosure allows to concentrate to high concentration the following elements, within the hydrolyser: scandium (Sc), galium (Ga), yttrium (Y), dysperosium (Dy), cerium (Ce), praseodynium (Pr), neodynium (Nd), europium (Eu), lanthanum (La), samarium (Sm), gadolinium, (Gd), erbium (Er), zirconium (Zr) and mixtures of thereof. Technologies that can be used for extracting rare earth elements and/or rare metals can be found, for example, in Zhou et al. in RARE METALS, Vol. 27, No. 3, 2008, p 223-227, and in US 2004/0042945, hereby incorporated by reference in their entirety. The person skilled in the art will also understand that various other processes normally used for extracting rare earth elements and/or rare metals from the Bayer process can also be used. For example, various solvent extraction techniques can be used. For certain elements, a technique involving octyiphenyl acid phosphate (OPAP) and toluene can be used. HCl can be used as a stripping agent. This can be effective for recovering Ce2O3, Sc2O3, Er2O3 etc. For example, different sequence using oxalic acid and metallic iron for ferric chloride separation can be used.


The spent acid liquor from steps 6 and 10 rich in value added metals, mainly magnesium, is processed to step 7. The solution is saturated with dry and highly concentrated gaseous hydrogen chloride from step 5, which results in the precipitation of magnesium chloride hexahydrate. For example, same can be accomplished with HCl in aqueous form over 30% wt. The precipitate retained, is fed to a calcination stage step 8 where pure MgO (>98% wt.) is obtained and highly concentrated hydrochloric acid (for example of at least 38%) is regenerated and diverted to the leaching step (step 2). An alternative route for step 7 is using dry gaseous hydrochloric acid from step 8.


In step 9, metal chlorides unconverted are processed to a pyrohydrolysis step (700-900° C.) to generate mixed oxides and where hydrochloric acid from 15-20.2% wt. concentration can be recovered.


According to another example as shown in FIG. 3, the processes can be similar to the example shown in FIG. 1 but can comprise some variants as below discussed.


In fact, as shown in FIG. 3, the processes can comprise (after step 6 or just before step 10) an internal recirculation back to the crystallization step 4. In such a case, The mother liquor from the hydrolyser (step 6) can be recirculated fully or partially to the crystallization of step 4 where a concentration increase will occur with respect to the non-hydrolysable elements including rare earth elements and/or rare metals.


Such a step can be useful for significantly increasing the concentration of rare earth elements and/or rare metals, thereby facilitating their extraction in step 10.


With respect to step 7, the solution rich in magnesium chloride and other non-hydrolysable products at low temperature is, as previously discussed, then brought up in concentration with dry and highly concentrated gaseous hydrogen chloride by sparging it into a crystallizer. This can result into the precipitation of magnesium chloride as an hexahydrate (for example after sodium and potassium chloride removal). This can also be accomplished with HCl in aqueous form.


As shown in FIG. 3, an extra step 11 can be added. Sodium chloride can undergo a chemical reaction with sulfuric acid so as to obtain sodium sulfate and regenerate hydrochloric acid at a concentration at or above the azeotropic point. Potassium chloride can undergo a chemical reaction with sulfuric acid so as to obtain potassium sulfate and regenerate hydrochloric acid at a concentration above the azeotropic concentration. Sodium and potassium chloride brine solution can be the feed material to adapted small chlor-alkali electrolysis cells. In this latter case, common bases (NaOH and KOH) and bleach (NaOCl and KOCl) are produced and can be reused to some extent in other areas of the processes of the present disclosure (scrubber, etc.).


The following are non-limitative examples.


Example 1
Preparation of Alumina and Various Other Products

As a starting material a sample of clay was obtained from the Grande Vallée area in Québec, Canada.


These results represent an average of 80 tests carried out from samples of about 900 kg each.


Crude clay in the freshly mined state after grinding and classification had the following composition:


Al2O3: 15%-26%;


SiO2: 45%-50%;


Fe2O3: 8%-9%;


MgO: 1%-2%;


Rare earth elements and/or rare metals: 0.04%-0.07%;


LOI: 5%-10%.


This material is thereafter leached in a two-stage procedure at 140-170° C. with 18-32 weight % HCl. The HCl solution was used in a stoichiometric excess of 10-20% based on the stoichiometric quantity required for the removal of the acid leachable constituents of the clay. In the first leaching stage of the semi-continuous operation (step 2), the clay was contacted for 2.5 hours with required amount or certain proportion of the total amount of hydrochloric acid. After removal of the spent acid, the clay was contacted again with a minimum 18 weight % hydrochloric acid solution for about 1.5 hour at same temperature and pressure.


A typical extraction curve obtained for both iron and aluminum for a single stage leaching is shown in FIG. 2.


The leachate was filtered and the solid was washed with water and analyzed using conventional analysis techniques (see step 3 of FIG. 1). Purity of obtained silica was of 95.4% and it was free of any chlorides and of HCl.


In another example, the purity of the silica was 99.67% through an extra leaching step.


After the leaching and silica removal, the concentration of the various metal chlorides was:


AlCl3: 15-20%;


FeCl2: 4-6%;


FeCl3: 0.5-2.0%;


MgCl2: 0.5-2.0%;


REE-Cl: 0.1-2%


Free HCl: 5-50 g/l


Spent acid was then crystallized using about 90 to about 98% pure dry hydrochloric acid in gas phase in two stages with less than 25 ppm iron in the aluminum chloride hexahydrate formed. The concentration of HCl in solution (aqueous phase) was about 22 to about 32% or 25 to about 32%, allowing 95.3% of Al2O3 recovery. The recovered crystallized material (hydrate form of AlCl3 having a minimum purity of 99.8%) was then calcined at 930° C. or 1250° C., thus obtaining the α form of the alumina. Heating at 930° C. allows for obtaining the beta-form of alumina while heating at 1250° C. allows for obtaining the alpha-form.


Another example was carried out at low temperature (decomposition and calcination at about 350° C.) and the α form of the alumina was less than 2%.


HCl concentration in gas phase exiting the calcination stage was having a concentration greater than 30% and was used (recycled) for crystallization of the AlCl3 and MgCl2. Excess of hydrochloric acid is absorbed at the required and targeted concentration for the leaching steps.


Iron chloride (about 90-95% in ferric form) is then sent to a hydrothermal process in view of its extraction as pure hematite (Fe2O3). This can be done by using the technology described in WO 2009/153321 of low temperature hydrolysis with full heat recovery from calcining, pyrohydrolysis and leaching stage.


Rare earth elements and rare metals are extracted from the mother liquor of the hydrolyzer where silica, aluminum, iron and a great portion of water have been removed and following preconcentration from hydrolyser to crystallization. It was observed that rare earth elements can be concentrated by a factor of about 4.0 to 10.0 on average within the hydrolyzer itself on a single pass through it i.e. without concentration loop. The following concentration factors have been noted within the hydrolyzer (single pass):

    • Ce>6
    • La>9
    • Nd>7
    • Y>9


Remaining magnesium chloride is sparged with dry and highly concentrated hydrochloric acid and then calcinated to MgO while recovering high concentration acid (for example up to 38.4%).


Mixed oxides (Me-O) containing other non-hydrolysable components were then undergoing a pyrohydrolysis reaction at 700-800° C. and recovered acid (15-20.2% wt.) was rerouted for example to the leaching system.


Overall Yields Obtained:


Al2O3: 93.0-95.03% recovery;


Fe2O3: 92.65-99.5% recovery;


Rare earth elements: 95% minimum recovery (mixture);


MgO: 92.64-98.00% recovery;


Material discarded: 0-5% maximum;


HCl global recovery: 99.75% minimum;


HCl strength as feed to leaching 15-32% (aqueous); 95% (gas)


Red mud production: none.


Example 2
Preparation of Alumina and Various Other Products

A similar feed material (bauxite instead of clay) was processed as per in example 1 up to the leaching stage and revealed to be easily leachable under the conditions established in example 1. It provided an extraction percentage of 100% for the iron and over 90-95% for aluminum. The technology was found to be economically viable and no harmful by-products (red mud) were generated. Samples tested had various concentrations of Al2O3 (up to 51%), Fe2O3 (up to 27%) and MgO (up to 1.5%). Gallium extraction of 97.0% was observed. Scandium extraction was 95%.


Example 3
HCl Gas Enrichment and Purification
H2SO4 Route

H2SO4 can be used for carrying out purification of HCl. It can be carried out by using a packing column with H2SO4 flowing counter currently (see FIG. 4). This allows for converting the recovered HCl into HCl having a concentration above the azeotropic point (20.1% wt) and increase its concentration by about 60 to about 70% at minimum.


Water is absorbed by H2SO4 and then H2SO4 regeneration is applied where H2SO4 is brought back to a concentration of about 95 to about 98% wt. Water release at this stage free of sulphur is recycled back and used for crystallization dissolution, etc. Packing of the column can comprise polypropylene or polytrimethylene terephthalate (PTT).


Combustion energy can be performed with off gas preheating air and oxygen enrichment. Oxygen enrichment: +2% represents flame temperature increase by: 400° C. maximum.


Thus, HCl of the processes of the present disclosure can thus be treated accordingly.


Example 4
HCl Gas Enrichment and Purification
Calcium Chloride to Calcium Chloride Hexahydrate (Absorption/Desorption Process)

As shown in FIG. 5, CaCl2 can be used for drying HCl. In fact, CaCl2 can be used for absorbing water contained into HCl. In such a case, CaCl2 is converted into its hexachloride form (CaCl2.6H2O) and one saturated system is eventually switched into regeneration mode where hot air recovered from calcination off gas of alumina and magnesium oxide spray roasting is introduced to regenerate the fixed bed. Alternatively, other absorbing agent such as LiCl can be used instead of CaCl2. Such an ion/exchange type process can be seen in FIG. 4 and the cycle can be inversed to switch from one column to another one.


The person skilled in the art would understand that the processes described in examples 3 and 4 (see FIGS. 4 and 5) can be used in various different manners. For example, these processes can be combined with the various processes presented in the present disclosure. For example, such purifications techniques can be integrated to the processes shown in FIG. 1, 3, 6, to 8 or 11 to 14. For example, these techniques can be used downstream of at least one of step chosen from steps 5, 6, 8, 9, 10, 11, 13 and 20 (see FIGS. 1, 3, 8, 13 and 14). They can also be used downstream of step 4 and/or step 7. They can also be used downstream of at least one of step chosen from steps 104 to 111 (see FIGS. 6 and 7). Moreover, they can be used in FIGS. 11 and 12 for example in steps 215, 216, 315 or 316.


Example 5
Preparation of Alumina and Various Other Products

This example was carried out by using a process as represented in FIGS. 6 and 7. It should be noted that the processes represented in FIGS. 6 and 7 differ mainly by the fact that FIG. 7 shows an stage i.e. stage 112.


Raw Material Preparation


Raw material, clay for example, was processed in a secondary crusher in the clay preparation plant 101. Dry milling and classifying occurs on a dry basis in vertical roller mills (for example Fuller-Loesche LM 30.41). The clay preparation 101 included three roller mills; two running at a capacity of approximately 160-180 tph and one on standby. Raw material, if required, can be reduced to 85% less than 63 microns. Processed material was then stored in homogenization silos before being fed to the acid leaching plant 102. Below in Table 1 are shown results obtained during stage 101. If the ore contains the fluorine element, a special treatment can be applied before carrying out the 102 stage. In presence of hydrochloric acid, fluorine can produce hydrofluoric acid. This acid is extremely corrosive and damaging for human health. Thus, before leaching 102, an optional treatment fluorine separation 112 can be done. Stage 112 can comprise treating the processed material coming from stage 101 with an acid in a pre-leaching treatment so as to remove hydrofluoric acid. Therefore, depending on the composition of the raw material, a fluorine separation stage 112 (or pre-leaching stage 112) can be carried out.









TABLE 1





Clay preparation


















Rate
290 tph











Composition feed
SiO2:
50.9%



(main constituents)
Al2O3:
24.0%




Fe2O3:
8.51%




CaO:
0.48%




MgO:
1.33%




Na2O:
1.06%




K2O:
2.86%




MnO:
0.16%




Cr2O3:
0.01%




TiO2:
0.85%




P2O5:
0.145% 




SrO:
0.015% 




BaO:
0.05%




V2O5
0.0321% 




Other (including
9.63%




H2O and REE):











Obtained particle size
85% < 63 μm



Residual moisture
0.5-0.7%



Yield
99.5% min










Acid Leaching


Next, acid leaching 102 was performed semi-continuously in an 80 m3 glass-lined reactor. Semi-continuous mode comprises replacing reacted acid ⅓ in the reaction period with higher concentration regenerated acid, which greatly improves reaction kinetics. The reactor arrangement comprises for example, a series of three reactors. Other examples have been carried out with a first leaching at 1 atm was carried out and then, a second and third semi-continuous or continuous leaching was carried out with aqueous or gaseous HCl.


Leaching was performed at high temperature and pressure (about 160 to about 195° C. and pressures of about 5 to about 8 barg) for a fixed period of time. Reaction time was a function of the reaction extent targeted (98% for Al2O3), leaching mode, acid strength, and temperature/pressure applied.


Spent acid recovered out of the acid leaching 102 was then filtered 103 from unreacted silica and titanium dioxide and washed through an automated filter press where all free HCl and chloride are recovered. Step 113 can then be carried out in various manners as indicated previously for step 13. This allows, for example, a maximum quantity of about 30 ppm SiO2 going into spent liquor. Cleaned silica at a concentration of ≈4196%+SiO2 is then produced. Various options are possible at that point. For example, the 96% silica can undergo final neutralization through caustic bath, cleaning, and then bricketing before storage. According to another example, the silica purified by adding another leaching step followed by a solid separation step that ensures TiO2 removal (see stage 113 in FIGS. 6 and 7). In that specific case, high purity silica 99.5%+ is produced. In stage 113, titanium and silicium can be separated from one another in various manners. For example, the solid obtained from stage 103 can be leached in the presence of MgCl2 at a temperature below 90 or 80° C. and at low acid concentration. For example, acid concentration can be below 25 or 20%. The acid can be HCl or H2SO4. In such a case, titanium remains soluble after such a leaching while titanium is still in a solid form. The same also applies when the solid is treated with Cl2. These solid and liquid obtained after stage 113 are thus separated to provide eventually TiO2 and SiO2. Water input and flow for silica cleaning is in a ratio of 1:1 (silica/water) (150 t/h SiO2/150 t/h H2O), but comprises of wash water circulation in closed loop in the process and limited amount of process water for final cleaning of the silica and recovery of all chlorides and free HCl generated at the leaching stage. Below in Table 2 are shown results obtained during stage 102.









TABLE 2





Acid Leaching
















Equivalent solid feed rate
259.6 tph


Operation mode
Semi-continuous


Acid to clay ratio
3.10 @ 23% wt



(Equivalent to 3.35 with semi-continuous at



18.0% wt)


Regenerated acid
18.0-32.0%


concentration



Operating temperature
150-155° C. (Pilot)



165-200° C. (Plant)


MAWP
120 psig


Typical chemical
Fe2O3 + 6 HCl → 2 FeCl3 + 3H2O


reactions
Al2O3 + 6 HCl → 2 AlCl3 + 3 H2O



MgO + 2 HCl → MgCl2 + H2O



K2O + 2 HCl → 2 KCl + H2O



Re2O3 + 6 HCl → 2 ReCl3 + 3H2O


Spent acid flow to
600-1100 m3/h


crystallization










Practical chemical
FeCl3
4.33%


composition after step
FeCl2
0.19%


102 without solid (SiO2)
AlCl3
16.6%



MgCl2
0.82%



NaCl
 1.1%



KCl
 1.2%



CaCl2
0.26%


Extraction yields
Iron
 100%



Al2O3

98%



SiO2 Recovery
99.997%









Energy consumption
Activation energy only and self-sustained



exothermic reaction from 130° C.










AlCl3 Crystallization


Spent acid, with an aluminum chloride content of about 20 to about 30%, was then processed in the crystallization stage 104. Dry and highly concentrated HCl (>90% wt.) in gas phase was sparged in a two-stage crystallization reactor, which allows the crystallization of aluminum chloride hexahydrate.


The flow rate of acid through these reactors is about 600 to about 675 m3/h and the reactor was maintained at about 50 to about 60° C. during this highly exothermic reaction. Heat was recovered and exchanged to the acid purification 107 part of the plant thus ensuring proper heat transfer and minimizing heat consumption of the plant. Aluminum chloride solubility decreases rapidly, compared to other elements, with the increase in concentration of free HCl in the crystallization reactor. The concentration of AlCl3 for precipitation/crystallization was about 30%


The HCl concentration during crystallization was thus about 30 to about 32% wt.


The aqueous solution from the crystallization stage 104 was then submitted to the hydrothermal acid recovery plant 105, while the crystals are processed through the decomposition/calcination stage in the calcination plant 106.


A one-step crystallization stage or a multi-step crystallization stage can be done. For example, a two-steps crystallization stage can be carried out.


Below in Tables 3A and 3B are shown results obtained during stage 104.









TABLE 3A





Aluminum chloride crystallization


















Number of crystallization steps
2



Operating temperature
50-60° C.



Sparging HCl concentration
90% (gaseous)



Typical chemicals formed
AlCl3•6H2O (s)




Metal chlorides (aq)



AlCl3•6H2O residual
<5% (practical); 8%
















TABLE 3B







Typical crystals composition main constituents


obtained at pilot scale and feeding calcination










Component
Weight distribution (%)













AlCl3•6H2O
99.978



BaCl2•2H2O
0.0000



CaCl2•6H2O
0.0009



CrCl4
0.0022



CuCl2•2H2O
0.0000



FeCl3•6H2O
0.0019



KCl
0.0063



MgCl2•6H2O
0.0093



MnCl2•4H2O
0.0011



NaCl
0.0021



SiCl4
0.0004



SrCl2•6H2O
0.0000



TiCl4
0.0001



VCl4
0.0000



Free Cl
0.0000










Calcination and Hydrothermal Acid Recovery


The calcination 106 comprises the use of a two-stage circulating fluid bed (CFB) with preheating systems. The preheating system can comprise a plasma torch to heat up steam to process. It processes crystals in the decomposition/calcination stage. The majority of the hydrochloric acid was released in the first stage which was operated at a temperature of about 350° C., while the second stage performs the calcination itself. Acid from both stages (about 66 to about 68% of the recovered acid from the processes) was then recovered and sent to either to the acid leaching 102 or to the acid purification 107. In the second reactor, which was operated at a temperature of about 930° C., acid was recovered through the condensation and absorption into two columns using mainly wash water from the acid leaching sector 102. Latent heat from this sector was recovered at the same time as large amounts of water, which limits net water input.


In the iron oxides productions and acid recovery 105 system, which comprises, aqueous solution from the crystallization 104 first undergoes a pre-concentration stage followed by processing in the hydrolyzer reactor. Here, hematite was produced during low temperature processing (about 165° C.). A recirculation loop was then taken from the hydrolyzer and is recirculated to the pre-concentrator, allowing the concentration of REE, Mg, K, and other elements. This recirculation loop, allows rare earth element chlorides and/or rare metal chlorides and various metal chlorides concentration to increase without having these products precipitating with hematite up to a certain extent.


Depending on acid balance in the plant, recovered acid is sent either directly to the 102 or 107 stage. Table 4 shows results obtained in stage 105.









TABLE 4







Hydrothermal acid recovery










Flowrate from crystallization to HARP
592 m3/h (design)




600 m3/h (design)



Operating hydrolyser temperature
155-170° C.



Regenerated acid concentration
27.4%



Regenerated acid flowrate
205.2 tph HCl



Hematite total production rate
24 TPH (design)



HCl recovery
>99.8%



Reflux (recirculation loop) rate in between
56 tph



hydrolyzer and preconcentrator




Rare earth element chlorides and/or rare
≈12.8 t/h



metal chlorides rate in recirculation loop










Hematite quality obtained and/or projected










Fe2O3 purity
>99.5%



Hydrolysable chlorides
<0.2%



Moisture
Max 20% after filtration



PSD
25-35 microns



Density (bulk)
2-3 kg/l










Typical chemical reaction in stage 105


2FeCl3 + 3H2O → Fe2O3 + 6 HCl


155-170° C.









Table 5 shows results obtained in stage 106.









TABLE 5







Calcination Plant 106








Process characteristics:
Two-stage circulating fluid bed



(CFB) with pre-heating system



Two-stage hydrochloric acid regeneration


Production rate (practical)
About 66 tph


CFB feed rate
371 tph @ 2-3% humidity*










Typical chemical reaction occurring


2(AlCl3•6 H2O) + Energy → Al2O3 + 6 HCl + 9H2O





Typical alumina chemical composition obtained from aluminum


chloride hexahydrate crystals being fed to calcination










Component
Weight distribution (%)






Al2O3
99.938



Fe2O3
0.0033



SiO2
0.0032



Cr2O3
0.0063



V2O5
0.0077



Na
0.0190



MgO
0.0090



P2O5
0.0039



K
0.0053



Ca
0.0020



MnO
0.0002



Free Cl
Undetectable










Rare Earth Elements and Rare Metals Extractions


The stream that was taken out of 105 recirculation then was treated for rare earth elements and are metals extraction 108, in which the reduction of the remaining iron back to iron 2 (Fe2+), followed by a series of solvent extraction stages, was performed. The reactants were oxalic acid, NaOH, DEHPA (Di-(2-ethylhexyl)phosphoric acid) and TBP (tri-n-butyl phosphate) organic solution, kerosene, and HCl were used to convert rare earth element chlorides and rare metals chlorides to hydroxides. Countercurrent organic solvent with stripping of solution using HCl before proceeding to specific calcination from the rare earth elements and rare metals in form of hydroxide and conversion to high purity individual oxides. A ion exchange technique is also capable of achieving same results as polytrimethylen terephtalate (PET) membrane.


Iron powder from 105, or scrap metal as FeO, can be used at a rate dependent on Fe3+ concentration in the mother liquor. HCl (100% wt) at the rate of 1 tph can be required as the stripped solution in REE Solvent Extraction (SX) separation and re-leaching of rare earth elements and/or rare metals oxalates.


Water of very high quality, demineralized or nano, at the rate of 100 tph was added to the strip solution and washing of precipitates.


Oxalic acid as di-hydrate at a rate of 0.2 tph was added and contributes to the rare earth elements and rare metals oxalates precipitation. NaOH or MgOH at a rate of 0.5 tph can be used as a neutralization agent.


DEHPA SX organic solution at the rate of 500 g/h was used as active reagent in rare earth elements separation while TBP SX organic solution at the rate of 5 kg/h is used as the active reagent for gallium recovery and yttrium separation. Finally, a kerosene diluent was used at the rate of approximately 2 kg/h in all SX section. Calcination occurs in an electric rotary furnace via indirect heating to convert contents to REE2O3 (oxides form) and maintain product purity.


Results of various tests made regarding stage 108 are shown in Table 6.










TABLE 6








One line divided in subsections (5) to isolate the



following elements using solvent extraction:






Ga2O3



Y2O3



Sc2O3



Eu2O3 + Er2O3 + Dy2O3



Ce2O3 + Nd2O3 + Pr2O3













Equivalent output earths oxides
166.14 kg/h







Projected production as per pilot testing results












Incoming
Final extraction



Feed
(kg/h)
individual (kg/h)






Ga2O3
15.66
11.98



Sc2O3
9.06
8.11



Y2O3
22.56
20.22



La2O3
32.24
25.67



Ce2O3
61.37
51.82



Pr2O3
8.08
6.18



Nd2O3
30.3
27.24



Sm2O3
5.7
4.51



Eu2O3
1.06
0.95



Gd2O3
4.5
4.06



Dy2O3
3.9
3.55



Er2O3
2.1
1.86



Total
196.55
166.14









Global Yield: 84.53%

Alternatively, stage 108 can be carried out as described in WO/2012/126092 and/or WO/2012/149642, that are hereby incorporated by reference in their entirety.


The solution after stages 108 and 109 contained mainly MgCl2, NaCl, KCl, CaCl2, FeCl2/FeCl3, and AlCl3 (traces), and then undergoes the 111 stage. Na, K, Ca that follows the MgO can be extracted in stage 110 by crystallization in a specific order; Na first, followed by K, and then Ca. This technique can be employed for example in the Israeli Dead Sea salt processing plant to produce MgO and remove alkali from the raw material.


HCl Regeneration


Alkali (Na, K), once crystallized, was sent and processed in the alkali hydrochloric acid regeneration plant 110 for recovering highly concentrated hydrochloric acid (HCl). The process chosen for the conversion can generate value-added products


Various options are available to convert NaCl and KCl with intent of recovering HCl. One example can be to contact them with highly concentrated sulfuric acid (H2SO4), which generates sodium sulphate (Na2SO4) and potassium sulfate (K2SO4), respectively, and regenerates HCl at a concentration above 90% wt. Another example, is the use of a sodium and potassium chloride brine solution as the feed material to adapted small chlor-alkali electrolysis cells. In this latter case, common bases (NaOH and KOH) and bleach (NaOCl and KOCl) are produced. The electrolysis of both NaCl and KCl brine is done in different cells where the current is adjusted to meet the required chemical reaction. In both cases, it is a two-step process in which the brine is submitted to high current and base (NaOH or KOH) is produced with chlorine (Cl2) and hydrogen (H2). H2 and Cl2 are then submitted to a common flame where highly concentrated acid in gas (100% wt.) phase is produced and can be used directly in the crystallization stage 104, or to crystallization stages requiring dry highly concentrated acid.


Magnesium Oxide


The reduced flow, which was substantially free of most elements (for example AlCl3, FeCl3, REE-Cl, NaCl, KCl) and rich in MgCl2, was then submitted to the magnesium oxides plant 111. In the MgO, pyrohydrolysis of MgCl2 and any other leftover impurities were converted into oxide while regenerating acid. The first step was a pre-evaporator/crystallizer stage in which calcium is removed and converted into gypsum (CaSO4.2H2O) by a simple chemical reaction with sulfuric acid, for which separation of MgO is required. This increases the capacity of MgO roasting and also energy consumption slightly, while substantially recovering HCl. The next step was the specific pyrohydrolysis of MgO concentrated solution by spray roasting. Two (2) main products were generated; MgO that was further treated and HCl (about 18% wt.), which was either recycled back to the upstream leaching stage 102 or to the hydrochloric acid purification plant (107) The MgO-product derived from the spray roaster can require further washing, purification, and finally calcining depending on the quality targeted. The purification and calcining can comprise a washing-hydration step and standard calcining step.


The MgO from the spray roaster is highly chemically active and was directly charged into a water tank where it reacts with water to form magnesium hydroxide, which has poor solubility in water. The remaining traces of chlorides, like MgCl2, NaCl, dissolved in water. The Mg(OH)2 suspension, after settling in a thickener, was forwarded to vacuum drum filters, which remove the remaining water. The cleaned Mg(OH)2 is then forwarded into a calcination reactor where it is exposed to high temperatures in a vertical multi-stage furnace. Water from hydration is released and allows the transformation of the Mg(OH)2 to MgO and water. At this point, the magnesium oxide was of high purity (>99%).


HCl Purification


The hydrochloric acid purification stage 107 is effective for purifying HCl regenerated from different sectors (for example 105, 106, 111) and to increase its purity for crystallization, whereas dry highly concentrated acid (>90% wt.) can be used as the sparging agent. Stage 107 also allowed for controlling the concentration of the acid going back to stage 102 (about 22 to about 32% wt.) and allows total acid and water balance. Total plant water balance is performed mainly by reusing wash water as absorption medium, as quench agent or as dissolution medium at the crystallization stages.


For example, HCl purification can be carried out as shown in FIGS. 4 and 5.


For example, purification can be carried out by means of a membrane distillation process. The membrane distillation process applied here occurs when two aqueous liquids with different temperatures are separated through a hydrophobic membrane. The driving force of the process was supplied by the partial pressure vapour difference caused by the temperature gradient between these solutions. Vapour travels from the warm to the cold side. Without wishing to be bound to such a theory, the separation mechanism was based on the vapour/liquid equilibrium of the HCl/water liquid mixture. Practical application of such a technology has been applied to HCl/water, H2SO4/water systems and also on large commercial scales on aqueous solution of sodium chloride with the purpose of obtaining potable water from seawater and nano water production. Therefore membrane distillation was a separation process based on evaporation through a porous hydrophobic membrane. The process was performed at about 60° C. and was effective to recover heat from the 104 and 102 stage with an internal water circulation loop, in order to maintain a constant incoming temperature to the membranes. For example, eight membranes of 300,000 m2 equivalent surface area can be used per membrane to obtain a concentration of HCl well above the azeotropic point (i.e. >36%) of the ≈750 m3/h and final 90% concentration is then obtained through pressure distillation (rectification column).


Purification of HCl by processing thus regenerated acid through hydrophobic membrane and separating water from HCl; therefore increasing HCl concentration up to about 36% (above azeotropic point) and therefore allowing with a single stage of rectification through a pressure stripping column to obtain >90% in gaseous phase, for crystallization stage (sparging); and therefore controlling acid concentration into crystallization stages up to 30-35%(aq).


As indicated stage 107 was operated at about 60° C. and heat input provided by heat recovery from stages 102 to 110. Rectification column was operated at about 140° C. in the reboiler part. Net energy requirement was neutral (negative in fact at −3.5 Gj/t Al2O3) since both systems were in equilibrium and in balance.


For example, the acid purification can be carried out by using adsorption technology over an activated alumina bed. In continuous mode, at least two adsorption columns are required to achieve either adsorption in one of them and regeneration in the other one. Regeneration can be performed by feeding in counter-current a hot or depressurized gas. This technology will result in a purified gas at 100% wt.


For example, the acid purification can be made by using calcium chloride as entrainer of water. A lean hydrochloric acid solution is contacted with a strong calcium chloride solution through a column. The water is then removed from the hydrochloric acid solution and 99.9% gaseous HCl comes out of the process. Cooling water and cryogenic coolant is used to condense water traces in the HCl. The weak CaCl2 solution is concentrated by an evaporator that ensures the recuperation of calcium chloride. Depending on the impurities in the incoming HCl solution feed to the column, some metals can contaminate the calcium chloride concentrated solution. A precipitation with Ca(OH)2 and a filtration allows the removal of those impurities. The column can operate for example at 0.5 barg. This technology can allow for the recuperation of 98% of the HCl.









TABLE 7







shows the results obtained concerning the process shown in Fig. 6.























TOTAL


Composition
Stage 101
Stage 102
Stage 106
Stage 105
MgO
Stage 107
Stage 108
PRODUCED
















(% wt)
Yield (%)
Yeild (%)
Yeild (%)
Yeild (%)
tpy
Yeild (%)
Yeild (%)
Yeild (%)
Yeild (%)



















Main constituents











SiO2

99.997%






99.997%


Al

98.02%
95.03%





95.03%


Fe

100.00%

92.65%




92.65%


Mg

99.998%


29,756
92.64%


92.64%


Ca

99.998%






98.28%


Na

99.998%






92.76%


K

100.00%






93.97%


Others incl. H2O











RE/RM

99.80%

92.32%



84.67%
84.67%


By-Products











NaOH




68,556






NAOCl




9,269






KOH




73,211






KOCl




9,586






CaSO4




46,837






Reactants











H2SO4(*)




19,204






Fresh HCl M-UP






99.75%

99.75%


Total

98.55%
95.03%

256,419
92.64%
99.75%
84.67%










Tables 8 to 26 show results obtained concerning the products made in accordance with the process shown in FIG. 6 in comparison with standard of the industry.









TABLE 8







Chemical composition of obtained alumina













Standard used in



Element
% Weight*
industry














Al2O3
99.938
98.35 min



Fe2O3
0.0033
0.0100



SiO2
0.0032
0.0150



TiO2
0.0003
0.0030



V2O5
0.0008
0.0020



ZnO
0.0005
0.0030



Cr2O3
0.0003
N/A



MgO
0.0090
N/A



MnO
0.0002
N/A



P2O5
0.0039
0.0010



Cu
0.0030
N/A



Ca
0.0020
0.0030



Na
0.0190
0.4000



K
0.0053
0.0150



Li
0.0009
N/A



Ba
<0.00001
0.0000



Th
<0.000001
0.0000



U
<0.000001
0.0000



Free Cl
Not detectable
0.0000



LOI
<1.0000
<1.0000 









P2O5 removal technique can include, for example, after leaching, phosphorous precipitation using zirconium sulphate. It can be provided, for example, in a solution heated at 80 to about 90° C. or about 85 to about 95° C., under vacuum.









TABLE 9







Physical properties of obtained alumina













Standard used in



Property
Orbite Alumina
industry






PSD < 20 μm
 5-10%
N/A



PSD < 45 μm
10-12%
<10%



PSD > 75 μm
50-60%
N/A



SSA (m2/g)
60-85
60-80



Att. Index
10-12%
<10%



α Al2O3
 2-5%
<7-9% 
















TABLE 10







Chemical composition of obtained hematite










Element
% Weight













Fe2O3
>99.5%



Hydrolysable elements
<0.2%
















TABLE 11







Physical properties of obtained hematite*










Property
Orbite hematite














PSDmean
25-35
μm



Density (bulk)
2000-3000
kg/m3










Humidity after filtration
<10%





*Material can be produced as brickets













TABLE 12







Chemical composition of obtained silica










Element
% Weight






SiO2
>99.7



Al2O3
<0.25% 



MgO
≈0.1%



Fe2O3
≈0.1%



CaO
≈0.01% 



Na2O
<0.1%



K2O
<0.1%





Note:


Product may have unbleached cellulose fiber filter aid. Cellulose wood flour.













TABLE 13







Physical properties of obtained silica










Property
Orbite silica














PSDmean
10-20
μm



Specific surface area
34
m2/g



Density (bulk)
2000-2500
kg/m3










Humidity after filtration
<30%
















TABLE 14







Purity of obtained rare earth element oxides










Element
Purity (%)






Ga2O3
>99%



Sc2O3




Y2O3




La2O3




Ce2O3




Pr2O3




Nd2O3




Sm2O3




Eu2O3




Gd2O3




Dy2O3




Er2O3










Physical properties of obtained REE-O/RM-O










Property
Orbite REE-O/RM-O














PSDmean
2-30
μm



Density
5500-13000
kg/m3










LOI
<1%
















TABLE 15







Chemical composition of obtained MgO











Element
Typical
Specification















MgO
99.0+
98.35
min











CaO
0.0020
0.83












SiO2
0.0000
0.20
max



B2O3
0.0000
0.02
max



Al2O3
0.0300
0.12
max



Fe2O3
0.0160
0.57
max



MnO2
<0.14
0.14
max











LOI
0.7%
<1%
















TABLE 16







Physical properties of obtained MgO










Property
Orbite MgO






PSDmean
10 μm



Density
N/A



LOI
650 kg/m3
















TABLE 17







Chemical composition of obtained NaOH










Element
% Weight






Sodium hydroxide
32%



Water
68%
















TABLE 18







Physical properties of obtained NaOH










Property
Sodium hydroxide (NaOH)






Physical state
Liquid



Vapour pressure
14 mmHg



Viscosity
>1



Boiling point
100° C.



Melting point
 0° C.



Specific gravity
1.0
















TABLE 19







Chemical composition of obtained sodium hypochlorite (bleach)










Element
% Weight






Sodium hypochlorite
12%



Sodium hydroxide
<1%



Water
>80% 
















TABLE 20







Physical properties of obtained NaOCl










Property
Sodium hypochlorite (NaOCl)






Physical state
Liquid



Vapour pressure
1.6 kPa



Viscosity
N/A



Boiling point
100° C.



Melting point
 −3° C.



Specific gravity
1.2
















TABLE 21







Chemical composition of obtained potassium hydroxide










Element
% Weight






Potassium hydroxide
32%



Water
68%
















TABLE 22







Physical properties of obtained potassium hydroxide










Property
KOH






Physical state
Liquid



Vapour pressure
17.5 mmHg



Viscosity
N/A



Boiling point
100° C.



Melting point
N/A



Specific gravity
1.18
















TABLE 23







Chemical composition of obtained potassium hypochlorite (KOCl)










Element
% Weight






Potassium hypochlorite
12%



Potassium hydroxide
<1%



Water
>80% 
















TABLE 24







Physical properties of obtained potassium hypochlorite










Property
KOCl






Physical state
Liquid



Vapour pressure
N/A



Viscosity
N/A



Boiling point
103° C.



Melting point
N/A



Specific gravity
>1.0
















TABLE 25







Chemical composition of obtained calcium sulphate dihydrate










Element
% Weight






Calcium sulphate dihydrate
100%
















TABLE 26







Physical properties of obtained calcium sulphate dihydrate










Property
Orbite CaSO4•2H2O






Physical state
Solid



Specific gravity
2.32









In order to demonstrate the versatility of the processes of the present disclosure, several other tests have been made so as to shown that these processes can be applied to various sources of starting material.


Example 6

Another starting material has been used for preparing acidic compositions comprising various components. In fact, a material that is a concentrate of rare earth elements and rare metals (particularly rich in zirconium) has been tested. Table 27 shows the results carried out on such a starting material using a similar process as shown in FIGS. 1, 3, 6, 7, 13 and 14 and as detailed in Examples 1, 2 and 5. It can thus be inferred from the results shown in Table 27 that the various components present in the leaching (various metals such as aluminum, iron, magnesium as well as rare earth elements and rare metals) can be extracted from the obtained leaching composition and that they can eventually be isolated by the processes of the present disclosure such as, for example, those presented in Examples 1, 2 and 5.


Example 7

Other tests have been made in a similar manner as described in Example 6. In the present example, carbonatite has been used as a starting material. (see Table 28 below).









TABLE 27







Tests made on a zirconium rich material.












Composition
Average





measure and/
measured
Extraction
O All


Raw
or evaluated
for testing
rate measured
Orbite process


material
(% wt.)
(% wt .)
(ALP) (%)
recovery (%)














Al2O3
6.12
6.12
89.65
86.97


Fe2O3
15.80
15.80
99.50
97.51


SiO2
36.00
36.00
0.000
99.997


MgO
3.08
3.08
99.75
92.66


Na2O
1.13
1.13
99.50
99.50


K2O
2.12
2.12
99.50
99.50


CaO
6.10
6.10
99.50
99.00


S total
0.22
0.22

100.00


F
1.98
1.98
99.50
99.00


TiO2
0.13
0.13
0.000
99.03


V2O5
0.00
0.00
98.00
96.04


P2O5
1.10
1.10
98.00
96.04


MnO
0.43
0.43
98.00
96.04


ZrO2
12.43
12.43
22.70
20.43


Cr2O3
0.00
0.00
0.00
0.00


Ce2O3
3.05
3.045
97.31
92.98


La2O3
1.34
1.337
99.55
92.68


Nd2O3
1.55
1.551
98.40
94.79


Pr2O3
0.37
0.375
99.75
97.52


Sm2O3
0.15
0.151
88.75
84.80


Dy2O3
0.09
0.089
80.35
76.77


Er2O3
0.03
0.030
72.60
69.37


Eu2O3
0.03
0.027
85.57
81.76


Gd2O3
0.21
0.205
82.85
79.16


Ho2O3
0.01
0.013
77.10
73.67


Lu2O3
0.00
0.003
60.15
57.47


Tb2O3
0.02
0.022
78.05
74.58


Th
0.02
0.022
88.10
84.18


Tm2O3
0.00
0.004
66.85
63.88


U
0.01
0.014
81.90
78.26


Y2O3
0.30
0.300
72.70
69.46


Yb2O3
0.02
0.023
62.80
60.01


Ga2O3
0.02
0.016
96.90
92.59


Sc2O3
0.00
0.003
95.00
90.77


LOI (inc.
6.122023973
6.12




water)
















TABLE 28







Tests made on carbonatite












Composition
Average





measure and/
measured
Extraction
O All


Raw
or evaluated
for testing
rate measured
Orbite process


material
(% wt.)
(% wt.)
(ALP) (%)
recovery (%)














Al2O3
0.70
0.70
84.31
81.61


Fe2O3
11.22
11.22
94.14
92.15


SiO2
2.11
2.11
0.00003
99.997


MgO
6.50
6.500
100
96.25


Na2O
0.07
0.07
92.54
90.55


K2O
0.18
0.181
37.33
37.33


CaO
16.51
16.51
100
98.00


TiO2
0.00
0.000
0.00000
100.000


V2O5
0.00
0.000
0
100.000


P2O5
0.00
0.000
0
100.000


MnO
0.00
0.000
0
100.000


ZrO2
0.00
0.000
0
100.000


Cr2O3
0.00
0.000
0
100.000


Ce2O3
1.19
1.195
64.04
61.190


La2O3
0.46
0.463
63.86
61.018


Nd2O3
0.45
0.448
81.46
77.835


Pr2O3
0.14
0.142
67.59
64.582


Sm2O3
0.03
0.033
65.32
62.413


Dy2O3
0.00
0.000
78.12
74.644


Er2O3
0.00
0.000
86.15
82.316


Eu2O3
0.01
0.007
66.45
63.493


Gd2O3
0.01
0.013
54.46
52.037


Ho2O3
0.00
0.000
83.12
79.421


Lu2O3
0.00
0.000
88.86
84.906


Tb2O3
0.00
0.001
41.42
39.577


Th
0.06
0.065




Tm2O3
0.00
0.000
90.70
86.664


U
0.01
0.007




Y2O3
0.00
0.000
84.68
80.912


Yb2O3
0.00
0.000
85.11
81.323


Ga2O3
0.00
0.000
0
0.000


Sc2O3
0.00
0.000
0
0.000


LOI (inc.

60.33




water)









It can thus be inferred from the results shown in Table 28 that the various metals, rare earth elements and rare metals extracted present in the obtained leaching composition can eventually be isolated by the processes of the present disclosure such as, for example, those presented in Examples 1, 2 and 5.


The process shown in FIG. 8 is similar to the process of FIG. 1, with the exception that in FIG. 8, the term “aluminum” is replaced by a “first metal”. The person skilled in the art would thus understand that in accordance with the present disclosure, the processes can also encompass recovering various other products and using various types of material as starting material. The first metal can be chosen from Al, Fe, Ti, Zn, Ni, Co, Mg, Li, Mn, Cu, Au, Ag, Pd, Pt. and mixtures thereof etc. Such a process can thus be used for recovering various other metals than aluminum. Thus, the first metal will be precipitated as a chloride in stage 5 and eventually converted into an oxide.


In fact, the person skilled in the art would understand that by replacing in FIGS. 1, 3, 6 and 7 the term “aluminum” with the expression “first metal” the processes shown in these figures can be used to obtain various other products than alumina and also used for treating various different starting material. Thus, the first metal can be recovered as a chloride (as it is the case for aluminum chlorides in the processes of FIGS. 1, 3, 6, 7, 13 and 14) and all the other stages of these processes can thus be carried out (when applicable) depending on the nature of the starting material used.


In step 4, the first metal chloride can be precipitated or crystallized. In fact, the first metal can be removed from the leachate in various manner. For example, a precipitating agent can be added or HCl (for example gaseous) can be reacted with the liquid obtained from step 3 so as to cause precipitation and/or crytallization of the first metal chloride. Alternatively, the temperature of the leachate can be controlled so as to substantially selectively cause precipitation of the first metal chloride.


As previously indicated, the processes of the present disclosure can be efficient for treating material comprising Al, Fe, Ti, Zn, Ni, Co, Mg, Li, Mn, Cu, Au, Ag, Pd, Pt.


For example, when treating a material that comprises, for example, Mg and Fe, the material can be leached for example by using HCl. Then, while the mixture (comprising a solid and a liquid) so obtained is still hot, it can be treated so as to separate the solid from the solid (for example by means of a solid/liquid separation). That will be effective for removing solids such as Si and optionally others such as Ti. Thus, the liquid can be cooled down to a temperature of about 5 to about 70° C., about 10 to about 60° C., about 10 to about 50° C., about 10 to about 40° C., or about 15 to about 30° C. so as to substantially selectively precipitate or crystallize magnesium (for example as MgCl2 (first metal chloride in FIG. 8)), as shown in 4 of such a figure. Then, the first metal chloride can be converted as shown in 5 so as to obtain the first metal oxide. The iron can then be treated as in 6 of FIG. 8. The remainder of the process shown in FIG. 8 (stages 6 to 10) being as described previously for FIG. 1.


Other examples of processes for treating material comprising magnesium and iron can be as shown in FIGS. 13 and 14. The processes of FIGS. 13 and 14 are similar to the processes of FIGS. 1 and 3, respectively. The main differences reside in steps 20 and 21 of FIGS. 13 and 14.


In these two examples of FIGS. 13 and 14 aluminum is also treated. In fact, as it can be seen in stages 20 and 21 of FIGS. 13 and 14, Mg, Fe and Al, the material can be leached for example by using HCl. Then, while the mixture (comprising a solid and a liquid) so obtained is still hot, it can be treated so as to separate the solid from the liquid (for example by means of a solid/liquid separation (see stage 3)). That will be effective for removing solids such as Si and optionally others such as Ti. Thus, the liquid can be cooled down to a temperature of about 5 to about 70° C., about 10 to about 60° C., about 10 to about 50° C., about 10 to about 40° C., or about 15 to about 30° C. so as to substantially selectively precipitate or crystallize magnesium (for example as MgCl2 (see stage 20 in FIGS. 13 and 14)). Then, magnesium chloride can be converted into magnesium oxide as shown in 21 of FIGS. 13 and 14. HCl can then be recovered and treated as previously indicated The remainder of the process shown in FIG. 13 (stages 4 to 10) are as described previously for FIG. 1 and the remainder of the process shown in FIG. 14 (stages 4 to 10) are as described previously for FIG. 3.


As previously indicated, magnesium can be firstly removed from the leachate and then aluminum can be removed as shown in FIGS. 13 and 14. Alternatively, aluminum can be firstly removed from the leachate and then magnesium can be removed. In such a case, steps 20 and 21 of FIGS. 13 and 14 would be disposed between steps 4 and 6.


As another example, a mixture Ni/Co of a low concentration in the feed (0.5-2.0% wt) can be leached with HCl according to FIG. 9. For example, leaching can be carried out by using HCl having a concentration of about 18 to about 32 wt % in a first reactor then, by using HCl having concentration of about 90 to about 95% (gaseous) in a second reactor; and by optionally using HCl having concentration of about 90 to about 95% (gaseous) in an optional third reactor. Then selective crystallization with HCl bubbling, solubility of chlorides (cobalt chloride vs nickel chloride) is distinct based on HCl concentration. Hexahydrate chloride can then be processed (produced) and fed for example to standard spray roaster 600-640° C. or fluid bed in view of producing oxides. HCl can therefore be regenerated, sent to closed loop acid purification where it is dried. Excess HCl can also be absorbed at its isotropic point and be used to solvent extraction or leaching. For example, nickel or cobalt chloride can thus replace aluminum chloride in FIGS. 1, 3, 6, 7, 13 and 14. After the leaching, shown in FIG. 9, the leachate can thus be treated as the leachate described in the processes described in the present disclosure and those of FIGS. 1, 3, 6, 7, 13 and 14, with the exception that instead of aluminum chloride, nickel chloride or cobalt chloride will be treated.


A similar approach can be adopted when using a starting material that contains Mg and Li. Leaching can be carried out as shown in FIG. 9 and selective precipitation of LiCl over MgCl2 or selective precipitation of MgCl2 over LiCl by injecting HCl (for example gaseous HCl) can be done. Na can be removed by crystallisation first. K can then be removed by crystallisation Moreover, for further purification, LiCl and MgCl2 can be separated by difference of solubility and/or crystallization in water.


For example, platinum and palladium can also be treated similarly. Moreover, their separation can also be accomplished with ion exchange: selective crystallization in HCl is possible and can be temperature sensitive.



FIGS. 10A and 10B show methods for separating Si from Ti. For example, when using an ore as starting material, leaching can be carried out in the presence of Cl2 (optionally in the presence of carbon) so as to maintain Ti under the form of TiCl4 since in remains in solution (fluid) while Si remains solid (SiO2). Then, Ti (such as TiCl4) can be heated so as to be converted into TiO2. For example, it can be injected into a plasma torch for being purified.


Such a method for purifying Si and Ti can be used in all the processes of the present disclosure when there is a need for separating these two entities. See stage 13 in FIGS. 1, 3, 6, 7, 13 and 14 and stage 113 in FIG. 7.


The processes shown in FIGS. 11A, 11B, 12A and 12B are processes that can be useful for treating various materials that comprise, for example, Mg and other metals such as Ni and/or Co. These materials can also comprise other metals such as aluminum, iron etc. The processes of FIGS. 11A, 11B, 12A and 12B are similar, with the exception that magnesium remains in solution after step 204 in FIG. 11A, while magnesium is precipitated after step 304 in FIG. 12A.


Certain steps carried out in the processes of FIGS. 11A, 11B, 12A and 12 are similar to the steps of other processes described in the present disclosure.


For example, steps 201 and 301 are similar to step 101 of FIGS. 6 and 7. Moreover, steps 202 and 302 of FIGS. 11 and 12 are similar to step 102 of FIGS. 6 and 7.


Steps 203 and 303 of FIGS. 11 and 12 are similar to step 103 of FIGS. 6 and 7.


Steps 213 and 313 of FIGS. 11 and 12 are similar to step 113 of FIG. 7. With respect to steps 214 and 314, TiO2 can eventually be purified by means of a plasma torch.


Eventually, CaSO4.2H2O (gypsum) can be produced as detailed in steps 223 and 323. Finally, pursuant to steps 224, 324, 225 and 325 Na2SO4 and K2SO4 can be produced.


With respects to steps 213 and 313, TiO2 can be converted into TiCl2 and/or TiCl4 so as to solubilize the titanium. For example, this can be done by reacting TiO2 optionally with Cl2 and carbon (C) (see FIGS. 10A and 10B. Therefore, SiO2 and titanium can be separated from one another since SiO2 remains solid while titanium will be solubilized. For example, steps 213, 313, 214 and 314 can be carried out as detailed in FIG. 10.


Such processes are also efficient for achieving whole recovery of HCl.


Pursuant to Ni and/or Co precipitation (steps 212 and 312) LiOH can be precipitated and eventually washed in steps 208 and 308. Then, a further leaching can be carried out in steps 209 and 309 so as to extract further metals. For example, if the starting material to be used in the processes of FIGS. 11 and 12 contains aluminum, steps 210 and 310 can be carried out so as to precipitate AlCl3. Such a step (210 or 310) is similar to step 104 carried out in FIGS. 6 and 7. In an analogous manner, steps 205 and 305 of FIGS. 11 and 12 are similar to step 105 of FIGS. 6 and 7. Steps 206 and 306 of FIGS. 11 and 12 are similar to step 106 of FIGS. 6 and 7. HCl purification carried out in steps 215 and 315 is similar to step 107 carried out in FIGS. 6 and 7. As it can be seen in FIGS. 216 and 316, HCl is thus regenerated.


Alternatively, pursuant to step 209, and depending on the composition of the starting material used for the processes of FIGS. 11 and 12, steps 210 and 310 can be omitted or bypassed. Therefore, if substantially no aluminum is comprised within the starting material, or if the content in aluminum is considerably low after step 209, step 249 can be carried out. The same also applied to step 309 and 349 of FIG. 12. Then, pursuant to steps 249 and 349 of FIGS. 11 and 12 in which a mixture of various metal chlorides are obtained, calcination can be carried out in steps 217 and 317 so as to eventually obtain a mixture of various metal oxides.


Impurities obtained in steps 210 and 310 can be crystallized in steps 218 and 318. By doing so, NaCl (steps 219 and 319) and KCl (steps 221 and 321) can be crystallized. An electrolysis of NaCl (steps 220 and 320) and KCl (steps 222 and 322) can be carried out as previously indicated in the present disclosure.


Example 8

Tests have been made for treating a magnesium-containing material as starting material. The magnesium-containing material was serpentine (asbestos) obtained from Black Lake, Quebec, Canada. Tables 29 to 31 below shows results obtained when leaching such a material with HCl. The serpentine ore was leached with a 30% molar excess of HCl at a temperature of about 150 to about 160° C.









TABLE 29





Tests made on serpentine





















Asbestos 1
Asbestos 2
Asbestos 3
Asbestos 4






Mass In
880
800
1000
820



Mass Out
334
250
 325
323



% water
 35%
 45%
 45%
 45%



















Al
Fe
Na
K





Asbestos
Initial
%
1.38
3.62
0.25
0.34


1
compound
kg
12.144
31.856
2.2
2.992



Cake
%
1.87
1.11
0.33
0.3




kg
4.05977
2.40981
0.71643
0.6513



Yield
%
67%
92%
67%
78%



recovery







Asbestos
Initial
%
0.49
3.56
0.03
0.04


2
compound
kg
3.92
28.48
0.24
0.32



Cake
%
0.59
0.47
0.02
0.01




kg
0.81125
0.64625
0.0275
0.01375



Yield
%
79%
98%
89%
96%



recovery







Asbestos
Initial
%
0.58
4.13
0.16
0.08


3
compound
kg
5.8
41.3
1.6
0.8



Cake
%
0.06
0.44
0.01
0.01




kg
0.10725
0.7865
0.017875
0.017875



Yield
%
98%
98%
99%
98%



recovery







Asbestos
Initial
%
0.31
5.54
0.01
0.01


4
compound
kg
2.542
45.428
0.082
0.082



Cake
%
1.14
0.37
0.41
0.23




kg
2.02521
0.657305
0.728365
0.408595



Yield
%
20%
99%
−788%
−398%



recovery













Mg
Ca
Ti
Si





Asbestos
Initial
%
18.1
0.59
0.03
20.3


1
compound
kg
159.28
5.192
0.264
178.64



Cake
%
6.49
0.16
0.01
34.8




kg
14.08979
0.34736
0.02171
75.5508



Yield
%
91%
93%
92%
58%



recovery







Asbestos
Initial
%
23.5
0.13
0.01
16.7


2
compound
kg
188
1.04
0.08
133.6



Cake
%
2.88
0.05
0.007
38.2




kg
3.96
0.06875
0.009625
52.525



Yield
%
98%
93%
88%
61%



recovery







Asbestos
Initial
%
22.3
0.23
0.01
17.3


3
compound
kg
223
2.3
0.1
173



Cake
%
3.27
0.02
0.01
34.3




kg
5.845125
0.03575
0.017875
61.31125



Yield
%
97%
98%
82%
65%



recovery







Asbestos
Initial
%
22.9
0.03
0.01
15.4


4
compound
kg
187.78
0.246
0.082
126.28



Cake
%
2.5
0.2
0.005
37.1




kg
4.44125
0.3553
0.0088825
65.90815



Yield
%
98%
−44%
89%
48%



recovery





















TABLE 30







Chemical Composition of Serpentine











Concentration measured



Components
and/or evaluated (% wt.)






Al2O3
0.59-2.61



Fe2O3
5.09-7.92



SiO2
32.94-43.43



MgO
30.01-38.97



Na2O
 0.04-0.337



K2O
0.012-0.41 



CaO
0.04-0.83



TiO2
0.017-0.050



V2O5
0.00 



P2O5
0.00 



MnO
0.005-0.080



ZrO
0.0000



F
0.00 



Co
0.0000



Cr
0.076-0.101



Cd
0.0000



Zn
0.0000



Ni
0.0000



Cu
0.0000



Pb
0.0000



As
0.0000



Ga2O3
0.0000



Sc2O3
0.0000



Re2O3
 0.00000



LOI (inc. water)
15.0-20.0
















TABLE 31







Leaching of Serpentine - Recovery Yields










Components
Leaching extraction rate (%)













Al2O3
81.34



Fe2O3
96.70



SiO2
0.00003



MgO
96.01



Na2O
84.96



K2O
90.57



CaO
95.05



TiO2
0.00002



V2O5
0.00



P2O5
0.00



MnO
0.00



ZrO
0.00



F
0.00



Co
0.00



Cr
0.00



Cd
0.00



Zn
0.00



Ni
0.00



Cu
0.00



Pb
0.00



As
0.00









The results of Tables 29 to 31 thus show that the processes of FIGS. 11 to 14 can be carried out with success.


It was also observed that when obtaining such a leachate by leaching serpentine with HCl, it was possible to substantially selectively precipitate some metals by controlling certain parameters. In fact, it was found that magnesium chloride has a very low solubility as compared with other chlorides (such as AlCl3, FeCl3, CaCl2, NaCl, KCl, MnCl2, etc.), for example when the leachate is at a temperature of about 10 to about 60° C., about 10 to about 40° C., about 15 to about 30° C., about 15 to about 25° C. or about 20° C. (see FIGS. 16 and 17). Therefore, one possible way among others of removing magnesium chloride in a substantially selective manner was to leach serpentine and remove the unleached solid while the mixture of solid and leachate is still hot. Then, when the solid is removed, the leachate can be cooled down so as to substantially selectively precipitate magnesium chloride.


Moreover, it was observed, during tests made, that when the leachate has a concentration in HCl of about 16 to about 20%, about 17 to about 18%, or about 17.5% by weight, MgCl2 was selectively precipitated over FeCl3 (see FIG. 15). It was also observed that magnesium chloride can have a low solubility at a temperature of about 15 to about 30° C., about 15 to about 25° C. or about 20° C. (see FIGS. 16 and 17).


The process shown in FIG. 18 is similar to the process shown in FIG. 1. The main difference resides in the fact that in the process of FIG. 18 comprises stages 25 and 26 instead of stage 5 of FIG. 1. In fact, in FIG. 18, the process comprises, after crystallization of AlCl3, to convert AlCl3 into Al(OH)3 before calcining the latter product into Al2O3. For example conversion of AlCl3 into Al(OH)3 can be carried out by reacting AlCl3 with a base (for example KOH or NaOH). Calcination of Al(OH)3 into Al2O3 can be carried out at high temperature such as about 800 to about 1200° C. or about 1000 to about 1200° C.


In fact, in the processes and the methods of the present disclosure, calcination of AlCl3 can be replaced by calcination of Al(OH)3, as shown in FIG. 18 (see the differences between the processes of FIG. 1 and FIG. 18). For example, stages 25 and 26 can replace stage 5 of various processes and methods such as shown in FIG. 3, 8, 13 and 14 or stage 106 of FIGS. 6 and 7.


The processes of the present disclosure provide a plurality of important advantages and distinction over the known processes.


The processes of the present disclosure provide fully continuous and economical solutions that can successfully extract alumina from various type of materials while providing ultra pure secondary products of high added value including highly concentrated rare earth elements and rare metals. The technology described in the present disclosure allows for an innovative amount of total acid recovery and also for a ultra high concentration of recovered acid. When combing it to the fact that combined with a semi-continuous leaching approach that favors very high extraction yields and allows a specific method of crystallization of the aluminum chloride and concentration of other value added elements. These processes also allow for preparing aluminum with such a produced alumina.


Specifically through the type of equipment used (for example vertical roller mill) and its specific operation, raw material grinding, drying and classifying can be applicable to various kinds of material hardness (furnace slag for example), various types of humidity (up to 30%) and incoming particle sizes. The particle size established provides the advantage, at the leaching stage, of allowing optimal contact between the minerals and the acid and then allowing faster kinetics of reaction. Particles size employed reduces drastically the abrasion issue and allows for the use of a simplified metallurgy/lining when in contact with hydrochloric acid.


A further advantage of the processes of the present disclosure is the combined high temperature and high incoming hydrochloric acid concentration. Combined with a semi continuous operation where the free HCl driving force is used systematically, iron and aluminum extraction yields do respectively reach 100% and 98% in less than about 40 of the reference time of a basic batch process. Another advantage of higher HCl concentration than the concentration at azeotropic point is the potential of capacity increase. Again a higher HCl concentration than the concentration of HCl at the azeotropic point and the semi-continuous approach represent a substantial advance in the art. The same also applies for continuous leaching.


Another advantage in that technique used for the mother liquor separation from the silica after the leaching stage countercurrent wash, is that band filters provide ultra pure silica with expected purity exceeding 96%.


The crystallization of AlCl3 into AlCl3.6H2O using dried, cleaned and highly concentrated gaseous HCl as the sparging agent allows for a pure aluminum chloride hexahydrate with only few parts per million of iron and other impurities. A minimal number of stages are required to allow proper crystal growth.


The direct interconnection with the calcination of AlCl3.6H2O into Al2O3 which does produce very high concentration of gas allows the exact adjustment in continuous of the HCl concentration within the crystallizer and thus proper control of the crystal growth and crystallization process.


The applicants have now discovered fully integrated and continuous processes with substantially total hydrochloric acid recovery for the extraction of alumina and other value added products from various materials that contain aluminum (clay, bauxite, aluminosilicate materials, slag, red mud, fly ashes etc.) containing aluminum. In fact, the processes allows for the production of substantially pure alumina and other value added products purified such as purified silica, pure hematite, pure other minerals (ex: magnesium oxide) and rare earth elements products. In addition, the processes do not require thermal pre-treatment before the acid leach operation. Acid leach is carried out using semi-continuous techniques with high pressure and temperature conditions and very high regenerated hydrochloric acid concentration. In addition, the processes do not generate any residues not sellable, thus eliminating harmful residues to environment like in the case of alkaline processes.


The advantage of the high temperature calcination stage, in addition for allowing to control the α-form of alumina required, is effective for providing a concentration of hydrochloric acid in the aqueous form (>38%) that is higher than the concentration of HCl at the azeotropic point and thus providing a higher incoming HCl concentration to the leaching stage. The calcination stage hydrochloric acid network can be interconnected to two (2) crystallization systems and by pressure regulation excess HCl can be being absorbed at the highest possible aqueous concentration. The advantage of having a hexahydrate chloride with low moisture content (<2%) incoming feed allows for a continuous basis to recover acid at a concentration that is higher than the azeotropic concentration. This HCl balance and double usage into three (3) common parts of the processes and above azeotropic point is a substantial advance in the art.


Another advantage is the use of the incoming chemistry (ferric chloride) to the iron oxide and hydrochloric acid recovery unit where all excess heat load from any calcination part, pyrohydrolysis and leaching part is being recovered to preconcentrate the mother liquor in metal chloride, thus allowing, at very low temperature, the hydrolysis of the ferric chloride in the form of very pure hematite and the acid regeneration at the same concentration than at its azeotropic point.


A further major advantage of the instant process at the ferric chloride hydrolysis step is the possibility to concentrate rare earth elements in form of chlorides at very high concentration within the hydrolyser reactor through an internal loop between hydrolyzer and crystallization. The advantage in that the processes of the present disclosure benefit from the various steps where gradual concentration ratios are applied. Thus, at this stage, in addition to an internal concentration loop, having the silica, the aluminum, the iron and having in equilibrium a solution close to saturation (large amount of water evaporated, no presence of free hydrochloric acid) allows for taking rare earth elements and non-hydrolysable elements in parts per million into the incoming feed and to concentrate them in high percentage directly at the hydrolyser after ferric chloride removal Purification of the specific oxides (RE-O) can then be performed using various techniques when in percentage levels. The advantage is doubled here: concentration at very high level of rare earth elements using integrated process stages and most importantly the approach prevents from having the main stream (very diluted) of spent acid after the leaching step with the risk of contaminating the main aluminum chloride stream and thus affecting yields in Al2O3. Another important improvement of the art is that on top of being fully integrated, selective removal of components allows for the concentration of rare earth elements to relatively high concentration (percentages).


Another advantage of the process is again a selective crystallization of MgCl2 through the sparging of HCl from either the alumina calcination step or the magnesium oxide direct calcination where in both cases highly concentrated acid both in gaseous phase or in aqueous form are being generated. As previously indicated, Mg(OH)2 can also be obtained. As per aluminum chloride specific crystallization, the direct interconnection with the calcination reactor, the HCl gas very high concentration (about 85 to about 95%, about 90 to 95% or about 90% by weight) allows for exact adjustment in continuous of the crystallizer based on quality of magnesium oxide targeted. Should this process step (MgO production or other value added metal oxide) be required based on incoming process feed chemistry, the rare earth elements extraction point then be done after this additional step; the advantage being the extra concentration effect applied.


The pyrohydrolysis allows for the final conversion of any remaining chloride and the production of refined oxides that can be used (in case of clay as starting material) as a fertilizer and allowing the processing of large amount of wash water from the processes with the recovery hydrochloric acid in close loop at the azeotropic point for the leaching step. The advantage of this last step is related to the fact that it does totally close the process loop in terms of acid recovery and the insurance that no residues harmful to the environment are being generated while processing any type of raw material, as previously described.


A major contribution to the art is that the proposed fully integrated processes of the present disclosure is really allowing, among others, the processing of bauxite in an economic way while generating no red mud or harmful residues. In addition to the fact of being applicable to other natural of raw materials (any suitable aluminum-containing material or aluminous ores), the fact of using hydrochloric acid total recovery and a global concentration that is higher than the concentration at the azeotropic point (for example about 21% to about 38%), the selective extraction of value added secondary products and compliance (while remaining highly competitive on transformation cost) with environmental requirements, represent major advantages in the art.


It was thus demonstrated that the present disclosure provides fully integrated processes for the preparation of pure aluminum oxide using a hydrochloric acid treatment while producing high purity and high quality products (minerals) and extracting rare earth elements and rare metals.


With respect to the above-mentioned examples 1 to 5, the person skilled in the art will also understand that depending on the starting material used (for example, clays, argillite, bauxite, kaolin, serpentine, kyanite nepheline, aluminosilicate materials, mudstone, beryl, cryolite, garnet, spinel, niccolite, kamacite, taenite, limonite, garnierite, laterite, pentlandite, smithsonite, warikahnite, sphalerite, chalcopyrite, chalcocite, covellite, bornite, tetrahedrite, malachite, azurite, cuprite, chrysocolla, ecandrewsite, geikielite, pyrophanite, ilmenite, red mud, slag, fly ashes, industrial refractory materials etc.,) some parameters might need to be adjusted consequently. In fact, for example, certain parameters such as reaction time, concentration, temperature may vary in accordance with the reactivity of the selected starting material.


While a description was made with particular reference to the specific embodiments, it will be understood that numerous modifications thereto will appear to those skilled in the art. Accordingly, the above description and accompanying drawings should be taken as specific examples and not in a limiting sense.

Claims
  • 1. A process for preparing alumina, said process comprising: leaching an aluminum-containing material with HCl so as to obtain a leachate comprising aluminum ions, magnesium ions and a solid, and separating said solid from said leachate;substantially selectively precipitating MgCl2 from said leachate under conditions effective for controlling solubility of MgCl2 based on at least one parameter chosen from temperature, acid concentration and chlorides concentration and removing said MgCl2 from said leachate;reacting said leachate with HCl so as to obtain a liquid and a precipitate comprising said aluminum ions in the form of AlCl3, and separating said precipitate from said liquid;optionally reacting said precipitate with a base;treating said precipitate under conditions effective for converting said precipitate into Al2O3 and optionally recovering gaseous HCl so-produced; andtreating said MgCl2 under conditions effective for converting it into MgO and optionally recovering gaseous HCl so-produced,
  • 2. A process for preparing alumina, said process comprising: leaching an aluminum-containing material with HCl so as to obtain a leachate comprising aluminum ions, magnesium ions and a solid, and separating said solid from said leachate;substantially selectively precipitating MgCl2 from said leachate under conditions effective for controlling solubility of MgCl2 based on at least one parameter chosen from temperature, acid concentration and chlorides concentration and removing said MgCl2 from said leachate;reacting said leachate with HCl so as to obtain a liquid and a precipitate comprising said aluminum ions in the form of AlCl3, and separating said precipitate from said liquid;optionally reacting said precipitate with a base;treating said precipitate under conditions effective for converting said precipitate into Al2O3 and optionally recovering gaseous HCl so-produced; andtreating said MgCl2 under conditions effective for converting it into MgO and optionally recovering gaseous HCl so-produced,
  • 3. The process of claim 1, wherein said process is carried out without reacting said precipitate with said base.
  • 4. The process of claim 2, wherein said process is carried out without reacting said precipitate with said base.
  • 5. The process of claim 1, wherein said process is carried out by reacting said precipitate with said base.
  • 6. The process of claim 2, wherein said process is carried out by reacting said precipitate with said base.
  • 7. The process of claim 1, wherein said aluminum-containing material is leached with HCl having a concentration of about 25 to about 45 weight %.
  • 8. The process of claim 1, wherein said aluminum-containing material is leached with HCl having a concentration of about 25 to about 45 weight % at a temperature of about 160 to about 190° C.
  • 9. The process of claim 1, wherein said aluminum-containing material is leached with HCl having a concentration of about 18 to about 32 weight % at a temperature of about 160 to about 175° C.
  • 10. The process of claim 1, comprising calcining MgCl2 into MgO.
  • 11. The process of claim 1, comprising calcining MgCl2 into MgO and recycling the gaseous HCl so-produced by contacting it with water so as to obtain a composition having a concentration of about 25 to about 45 weight % and using said composition for leaching said aluminum-containing material.
  • 12. The process of claim 1, wherein said process comprises reacting said leachate with dry gaseous HCl so as to obtain said liquid and said precipitate comprising said aluminum ions, said precipitate being formed by crystallization of AlCl3.6H2O.
  • 13. The process of claim 1, wherein said process comprises reacting said leachate with HCl recovered during said process and having a concentration of at least 30% as to obtain said liquid and said precipitate comprising said aluminum ions, said precipitate being formed by crystallization of AlCl3.6H2O.
  • 14. The process of claim 1, wherein said crystallization is carried out at a temperature of about 45 to about 65° C.
  • 15. The process of claim 1, wherein said process comprises converting AlCl3.6H2O into Al2O3 by carrying out a calcination of AlCl3.6H2O, said calcination comprising steam injection.
  • 16. The process of claim 1, wherein said aluminum-containing material is chosen from aluminosilicate minerals.
  • 17. The process of claim 2, wherein MgCl2 is substantially selectively precipitated from said leachate and removed therefrom and then, said leachate is reacted with HCl so as to obtain said liquid and said precipitate comprising said aluminum ions in the form of AlCl3, and separating said precipitate from said liquid.
  • 18. The process of claim 2, wherein said leachate is reacted with HCl so as to obtain said liquid and said precipitate comprising said aluminum ions in the form of AlCl3, and separating said precipitate from said liquid, and then said MgCl2 is substantially selectively precipitated from said leachate and removed therefrom.
  • 19. The process of claim 1, wherein said aluminum-containing material is leached with HCl so as to obtain said leachate comprising aluminum ions, magnesium ions and said solid, and said solid is separated from said leachate at a temperature of at least 50° C.
  • 20. A process for preparing alumina, said process comprising: leaching an aluminum-containing material with HCl so as to obtain a leachate comprising aluminum ions, magnesium ions and iron ions and a solid, and separating said solid from said leachate;reacting said leachate with HCl so as to obtain a liquid comprising said magnesium ions and said iron ions and a precipitate comprising said aluminum ions in the form of AlCl3, and separating said precipitate from said liquid;substantially selectively removing said iron ions from said leachate by means of an hydrolysis;substantially selectively precipitating MgCl2 from said liquid under conditions effective for controlling solubility of MgCl2 based on at least one parameter chosen from temperature, acid concentration and chlorides concentration and removing said MgCl2 from said liquid;treating said precipitate under conditions effective for converting AlCl3 into Al2O3 and optionally recovering gaseous HCl so-produced; andtreating said MgCl2 under conditions effective for converting it into MgO and optionally recovering gaseous HCl so-produced.
CROSS-REFERENCE TO RELATED APPLICATIONS

The present application is a 35 USC 371 national stage entry of PCT/CA2013/000830 filed on Sep. 26, 2013 and which claims priority on U.S. 61/705,898 filed on Sep. 26, 2012, on U.S. 61/713,795 filed on Oct. 15, 2012; on U.S. 61/726,971 filed on Nov. 15, 2012; U.S. 61/837,715 filed on Jun. 21, 2013. These documents are hereby incorporated by reference in their entirety.

PCT Information
Filing Document Filing Date Country Kind
PCT/CA2013/000830 9/26/2013 WO 00
Publishing Document Publishing Date Country Kind
WO2014/047728 4/3/2014 WO A
US Referenced Citations (276)
Number Name Date Kind
558726 Gooch Apr 1896 A
650763 Raynaud May 1900 A
1494029 Scofield et al. May 1924 A
1501873 Tyrer Jul 1924 A
1519880 Heinrich et al. Dec 1924 A
1701510 Sieurin Feb 1929 A
1760962 Phillips et al. Jun 1930 A
1778083 Marburg Oct 1930 A
1906467 Heath May 1933 A
1931515 Fritz et al. Oct 1933 A
1956139 Staufer et al. Apr 1934 A
1962498 Frost Jun 1934 A
1999773 McMichael Apr 1935 A
2024026 Coleman et al. Dec 1935 A
2189376 Burman Feb 1940 A
2354133 Lyons Jul 1944 A
2376696 Hixson et al. May 1945 A
2406577 Alessandroni Aug 1946 A
2413709 Hoffman Jan 1947 A
2471844 Strelzoff May 1949 A
2489309 Mills et al. Nov 1949 A
2648595 Kennedy Aug 1953 A
2663620 Hinsdale, III Dec 1953 A
2707149 McKinley Apr 1955 A
2722471 Hirsch et al. Nov 1955 A
2769686 Michener, Jr. et al. Nov 1956 A
2771344 Michel et al. Nov 1956 A
2780525 Wendell, Jr. et al. Feb 1957 A
2806766 Anderson Sep 1957 A
2815264 Calkins et al. Dec 1957 A
2824783 Peppard et al. Feb 1958 A
2848398 Inagaki Aug 1958 A
2914381 Wainer Nov 1959 A
2914464 Burton et al. Nov 1959 A
2992893 Soudan et al. Jul 1961 A
3013859 Kuhlman, Jr. et al. Dec 1961 A
3104950 Ellis Sep 1963 A
3159452 Lerner Dec 1964 A
3192128 Brandmair et al. Jun 1965 A
3211521 George et al. Oct 1965 A
3473919 Metcalfe et al. Oct 1969 A
3479136 Michener, Jr. et al. Nov 1969 A
3540860 Cochran Nov 1970 A
3545920 George et al. Dec 1970 A
3586477 Flood Jun 1971 A
3620671 Maurel et al. Nov 1971 A
3642441 Van Weert Feb 1972 A
3649185 Sato et al. Mar 1972 A
3658483 Lienau et al. Apr 1972 A
3682592 Kovacs Aug 1972 A
3751553 Oslo et al. Aug 1973 A
3816605 Schwandorf Jun 1974 A
3852430 Lienau et al. Dec 1974 A
3862293 Maurel et al. Jan 1975 A
3903239 Berkovich Sep 1975 A
3922164 Reid et al. Nov 1975 A
3944648 Solymar et al. Mar 1976 A
3946103 Hund Mar 1976 A
3957504 Ho et al. May 1976 A
3966909 Grunig et al. Jun 1976 A
3983212 Lowenstein et al. Sep 1976 A
4042664 Cardwell et al. Aug 1977 A
4045537 Hrishikesan Aug 1977 A
4048285 Szepesi et al. Sep 1977 A
4069296 Huang Jan 1978 A
4098868 Tolley Jul 1978 A
4107281 Reh et al. Aug 1978 A
4110399 Gaudernack et al. Aug 1978 A
4124680 Cohen et al. Nov 1978 A
4130627 Russ et al. Dec 1978 A
4133677 Matsui et al. Jan 1979 A
4158042 Deutschman Jun 1979 A
4172879 Miller et al. Oct 1979 A
4177242 Cohen et al. Dec 1979 A
4193968 Sullivan et al. Mar 1980 A
4198231 Gusset Apr 1980 A
4222989 Belsky et al. Sep 1980 A
4224287 Ziegenbalg et al. Sep 1980 A
4226844 Reh et al. Oct 1980 A
4233273 Meyer et al. Nov 1980 A
4237102 Cohen et al. Dec 1980 A
4239735 Eisele et al. Dec 1980 A
4241030 Cohen et al. Dec 1980 A
4259311 Shah Mar 1981 A
4297326 Gjelsvik et al. Oct 1981 A
4318896 Schoonover Mar 1982 A
4362703 Boybay et al. Dec 1982 A
4370422 Panda et al. Jan 1983 A
4378275 Adamson et al. Mar 1983 A
4392987 Laine et al. Jul 1983 A
4402932 Miller et al. Sep 1983 A
4414196 Matsumoto et al. Nov 1983 A
4435365 Morris Mar 1984 A
4437994 Baker Mar 1984 A
4465566 Loutfy et al. Aug 1984 A
4465659 Cambridge et al. Aug 1984 A
4486393 Baksa et al. Dec 1984 A
4490338 De Schepper et al. Dec 1984 A
4530819 Czeglédi et al. Jul 1985 A
4560541 Davis Dec 1985 A
4567026 Liosowyj Jan 1986 A
4634581 Cambridge et al. Jan 1987 A
4652433 Ashworth et al. Mar 1987 A
4676838 Franz et al. Jun 1987 A
4710369 Bergman Dec 1987 A
4741831 Grinstead May 1988 A
4797271 Fleming et al. Jan 1989 A
4816233 Rourke et al. Mar 1989 A
4820498 Newkirk Apr 1989 A
4826671 Arndt et al. May 1989 A
4830507 Bagatto et al. May 1989 A
4898719 Rourke et al. Feb 1990 A
4913884 Feuling Apr 1990 A
4938871 Musikas et al. Jul 1990 A
4965053 Herchenroeder et al. Oct 1990 A
4968504 Rourke et al. Nov 1990 A
4980141 Kimura et al. Dec 1990 A
4988487 Lai et al. Jan 1991 A
4995984 Barkatt et al. Feb 1991 A
5006753 Hasker et al. Apr 1991 A
5008089 Moody et al. Apr 1991 A
5011665 Cailly et al. Apr 1991 A
5015447 Fulford et al. May 1991 A
5019362 Rourke et al. May 1991 A
5030424 Fulford et al. Jul 1991 A
5035365 Birmingham Jul 1991 A
5037608 Tarcy et al. Aug 1991 A
5039336 Feuling Aug 1991 A
5043077 Chandler et al. Aug 1991 A
5045209 Snyder et al. Sep 1991 A
5053144 Szirmai et al. Oct 1991 A
5061474 Pauli et al. Oct 1991 A
5071472 Traut et al. Dec 1991 A
5080803 Bagatto et al. Jan 1992 A
5091159 Connelly et al. Feb 1992 A
5091161 Harris et al. Feb 1992 A
5093091 Dauplaise et al. Mar 1992 A
5104544 Shimizu et al. Apr 1992 A
5106797 Allaire Apr 1992 A
5112534 Guon et al. May 1992 A
5120513 Moody et al. Jun 1992 A
5124008 Rendall et al. Jun 1992 A
5149412 Allaire Sep 1992 A
5160482 Ash et al. Nov 1992 A
5180563 Lai et al. Jan 1993 A
5188809 Crocker et al. Feb 1993 A
5192443 Delloye et al. Mar 1993 A
5244649 Ostertag et al. Sep 1993 A
5274129 Natale et al. Dec 1993 A
5368736 Horwitz et al. Nov 1994 A
5409677 Zinn Apr 1995 A
5409678 Smith et al. Apr 1995 A
5433931 Bosserman Jul 1995 A
5443618 Chapman Aug 1995 A
5492680 Odekirk Feb 1996 A
5500043 Harada et al. Mar 1996 A
5505857 Misra et al. Apr 1996 A
5512256 Bray et al. Apr 1996 A
5531970 Carlson Jul 1996 A
5585080 Andersen et al. Dec 1996 A
5597529 Tack Jan 1997 A
5622679 Yuan et al. Apr 1997 A
5632963 Schwab et al. May 1997 A
5639433 Yuan et al. Jun 1997 A
5645652 Okinaka et al. Jul 1997 A
5665244 Rothenberg et al. Sep 1997 A
5720882 Stendahl et al. Feb 1998 A
5723097 Barnett et al. Mar 1998 A
5766478 Smith et al. Jun 1998 A
5787332 Black et al. Jul 1998 A
5792330 Petersen et al. Aug 1998 A
5795482 Ehle et al. Aug 1998 A
5876584 Cortellini Mar 1999 A
5885545 Pitzer Mar 1999 A
5904856 Kvant et al. May 1999 A
5911967 Ruthner Jun 1999 A
5922403 Tecle Jul 1999 A
5942199 Jokinen et al. Aug 1999 A
5955042 Barnett et al. Sep 1999 A
5962125 Masaki Oct 1999 A
5993758 Nehari et al. Nov 1999 A
5997828 Rendall Dec 1999 A
6033579 Riemer et al. Mar 2000 A
6045631 Tarcy et al. Apr 2000 A
6077486 Spitzer Jun 2000 A
6093376 Moore Jul 2000 A
6153157 McLaughlin Nov 2000 A
6214306 Aubert et al. Apr 2001 B1
6221233 Rendall Apr 2001 B1
6238566 Yoshida et al. May 2001 B1
6248302 Barnett et al. Jun 2001 B1
6254782 Kreisler Jul 2001 B1
6267936 Delmas et al. Jul 2001 B1
6302952 Mobbs et al. Oct 2001 B1
6309441 Benz et al. Oct 2001 B1
6312653 Delmau et al. Nov 2001 B1
6337061 Iyatomi et al. Jan 2002 B1
6348154 Stewart Feb 2002 B1
6383255 Sundkvist May 2002 B1
6395062 Olafson et al. May 2002 B2
6395242 Allen et al. May 2002 B1
6406676 Sundkvist Jun 2002 B1
6447738 Rendall et al. Sep 2002 B1
6468483 Barnett et al. Oct 2002 B2
6500396 Lakshmanan et al. Dec 2002 B1
6565733 Sportel et al. May 2003 B1
6576204 Johansen Jun 2003 B2
6716353 Mirzadeh et al. Apr 2004 B1
6843970 Hard Jan 2005 B1
6893474 Jäfverström et al. May 2005 B2
7090809 Harel et al. Aug 2006 B2
7118719 Fugleberg Oct 2006 B2
7182931 Turnbaugh, Jr. et al. Feb 2007 B2
7220394 Sreeram et al. May 2007 B2
7282187 Brown et al. Oct 2007 B1
7294319 Lahtinen et al. Nov 2007 B2
7381690 Ding et al. Jun 2008 B1
7442361 Gloeckler et al. Oct 2008 B1
7498005 Yadav Mar 2009 B2
7651676 Beaulieu et al. Jan 2010 B2
7781365 Okamoto Aug 2010 B2
7837961 Boudreault et al. Nov 2010 B2
7892426 Hayashi et al. Feb 2011 B2
7906097 Beaulieu et al. Mar 2011 B2
8038969 Kondo et al. Oct 2011 B2
8147795 Dolling et al. Apr 2012 B2
8216532 Vierheilig Jul 2012 B1
8241594 Boudreault et al. Aug 2012 B2
8287826 Pettey Oct 2012 B2
8337789 Boudreault et al. Dec 2012 B2
8568671 Guo et al. Oct 2013 B2
8597600 Boudreault et al. Dec 2013 B2
20020014416 Van Weert Feb 2002 A1
20020050230 Meisen May 2002 A1
20020071802 Fulton et al. Jun 2002 A1
20030075021 Young et al. Apr 2003 A1
20030152502 Lewis et al. Aug 2003 A1
20030183043 Wai et al. Oct 2003 A1
20040042945 Rao et al. Mar 2004 A1
20040062695 Horwitz et al. Apr 2004 A1
20050166706 Withers et al. Aug 2005 A1
20060018813 Bray Jan 2006 A1
20060066998 Ishiguro Mar 2006 A1
20070062669 Song et al. Mar 2007 A1
20070278106 Shaw Dec 2007 A1
20080047395 Liu et al. Feb 2008 A1
20080069748 Lien et al. Mar 2008 A1
20080115627 Wang et al. May 2008 A1
20080286182 Costa et al. Nov 2008 A1
20090241731 Pereira et al. Oct 2009 A1
20090272230 Mackowski et al. Nov 2009 A1
20100018347 Holden et al. Jan 2010 A1
20100078382 Naganawa et al. Apr 2010 A1
20100129277 Kondo et al. May 2010 A1
20100150799 Boudreault et al. Jun 2010 A1
20100160144 Kim et al. Jun 2010 A1
20100260640 Shindo et al. Oct 2010 A1
20100278720 Wong et al. Nov 2010 A1
20100319491 Sugahara et al. Dec 2010 A1
20100329970 Lian et al. Dec 2010 A1
20110017020 Homma et al. Jan 2011 A1
20110044869 Boudreault et al. Feb 2011 A1
20110120267 Roche May 2011 A1
20110182786 Burba, III Jul 2011 A1
20120073407 Drinkard, Jr. et al. Mar 2012 A1
20120237418 Boudreault et al. Sep 2012 A1
20130052103 Boudreault et al. Feb 2013 A1
20130233130 Boudreault et al. Sep 2013 A1
20140065038 Boudreault et al. Mar 2014 A1
20140286841 Boudreault et al. Sep 2014 A1
20140301920 Boudreault et al. Oct 2014 A1
20140341790 Kasaini Nov 2014 A1
20140356262 Ruth et al. Dec 2014 A1
20140369904 Boudreault et al. Dec 2014 A1
20140369907 Boudreault et al. Dec 2014 A1
20140373683 Boudreault et al. Dec 2014 A1
Foreign Referenced Citations (169)
Number Date Country
631226 Feb 1991 AU
4375001 Dec 2001 AU
1066872 Nov 1979 CA
1088961 Nov 1980 CA
1136380 Nov 1982 CA
1176470 Oct 1984 CA
1224327 Jul 1987 CA
1226719 Sep 1987 CA
2027519 Apr 1991 CA
2027973 Apr 1991 CA
2029623 May 1991 CA
2036058 Aug 1991 CA
2097809 Jul 1992 CA
2137249 Dec 1993 CA
2122364 Feb 1994 CA
2156295 Sep 1994 CA
2160488 Nov 1994 CA
2193726 Jan 1996 CA
2159534 Apr 1996 CA
2167890 Jul 1996 CA
2240067 Jun 1997 CA
2251433 Apr 1999 CA
2360447 Aug 2000 CA
2306015 Dec 2000 CA
2309225 Dec 2000 CA
2377600 Jan 2001 CA
2317692 Mar 2001 CA
2391394 May 2001 CA
2400673 Aug 2001 CA
2429889 Jun 2002 CA
2431466 Jun 2002 CA
2433448 Jul 2002 CA
2189631 Nov 2002 CA
2454812 Feb 2003 CA
2468885 Jul 2003 CA
2471179 Jul 2003 CA
2378721 Sep 2003 CA
2484134 Nov 2003 CA
2467288 Nov 2004 CA
2548225 Nov 2004 CA
2385775 May 2005 CA
2556613 Aug 2005 CA
2572190 Jan 2006 CA
2597440 Aug 2006 CA
2521817 Mar 2007 CA
2624612 Apr 2007 CA
2629167 May 2007 CA
2639796 Jun 2007 CA
2636379 Jul 2007 CA
2641919 Aug 2007 CA
2538962 Sep 2007 CA
2608973 Jan 2008 CA
2610918 Feb 2008 CA
2659449 Feb 2008 CA
2684696 Nov 2008 CA
2685369 Nov 2008 CA
2711013 Nov 2008 CA
2697789 Mar 2009 CA
2725391 Nov 2009 CA
2678724 Mar 2010 CA
2745572 Jul 2010 CA
2747370 Jul 2010 CA
2667029 Nov 2010 CA
2667033 Nov 2010 CA
2678276 Mar 2011 CA
2773571 Mar 2011 CA
2788965 Aug 2011 CA
2797561 Nov 2011 CA
2834356 Dec 2012 CA
2884787 Apr 2013 CA
2860491 Aug 2013 CA
1099424 Mar 1995 CN
1923730 Mar 2007 CN
101773925 Jul 2010 CN
101792185 Aug 2010 CN
102849765 Jan 2013 CN
102849767 Jan 2013 CN
103420405 Dec 2013 CN
19903011 Aug 2000 DE
157503 Oct 1985 EP
0054976 Jul 1986 EP
0238185 Sep 1987 EP
0279672 Aug 1988 EP
0327234 Aug 1989 EP
0382383 Aug 1990 EP
0399786 Nov 1990 EP
508676 Oct 1992 EP
466338 Dec 1995 EP
0449942 Apr 1996 EP
0775753 May 1997 EP
0829454 Mar 1998 EP
0692035 Apr 1998 EP
0834584 Apr 1998 EP
999185 May 2000 EP
1496063 Jan 2005 EP
2241649 Oct 2010 EP
2298944 Mar 2011 EP
2319579 Feb 1977 FR
2600635 Dec 1987 FR
120035 Mar 1919 GB
153500 Nov 1920 GB
159086 Feb 1921 GB
195295 Mar 1923 GB
230916 Mar 1925 GB
240834 May 1926 GB
241184 May 1926 GB
273999 Jul 1927 GB
409710 May 1934 GB
470305 Aug 1937 GB
480921 Mar 1938 GB
490099 Aug 1938 GB
574818 Jan 1946 GB
745601 Feb 1956 GB
798750 Jul 1958 GB
857245 Dec 1960 GB
858026 Jan 1961 GB
1021326 Mar 1966 GB
1056488 Jan 1967 GB
1307319 Feb 1973 GB
2013164 Aug 1979 GB
1552918 Sep 1979 GB
2018230 Oct 1979 GB
2238813 Jun 1991 GB
05287405 Nov 1993 JP
6056429 Mar 1994 JP
010034 Oct 1996 OA
8603521 Jun 1986 WO
9103424 Mar 1991 WO
9213637 Aug 1992 WO
9313017 Jul 1993 WO
9418122 Aug 1994 WO
9600698 Jan 1996 WO
9624555 Aug 1996 WO
9722554 Jun 1997 WO
0017408 Mar 2000 WO
0104366 Jan 2001 WO
2004056468 Jul 2004 WO
2004056471 Jul 2004 WO
2004101833 Nov 2004 WO
2005123591 Dec 2005 WO
2006084682 Aug 2006 WO
2007074207 Jul 2007 WO
2007079532 Jul 2007 WO
2008067594 Jun 2008 WO
2008104250 Sep 2008 WO
2008141423 Nov 2008 WO
2008154995 Dec 2008 WO
2009085514 Jul 2009 WO
2009153321 Dec 2009 WO
2010002059 Jan 2010 WO
2010009512 Jan 2010 WO
2010056742 May 2010 WO
2010079369 Jul 2010 WO
2010133284 Nov 2010 WO
2011094858 Aug 2011 WO
2011100820 Aug 2011 WO
2011100821 Aug 2011 WO
2011147867 Dec 2011 WO
2012126092 Sep 2012 WO
2012145797 Nov 2012 WO
2012149642 Nov 2012 WO
2013037054 Mar 2013 WO
2013142957 Oct 2013 WO
2014029031 Feb 2014 WO
2014047728 Apr 2014 WO
2014075173 May 2014 WO
2014094155 Jun 2014 WO
2014094157 Jun 2014 WO
2014124539 Aug 2014 WO
Non-Patent Literature Citations (179)
Entry
Partial Translation of the Abstract and Claims for CN1099424A, “Method for treating blast furnace slag with diluted chlorhydric acid”, published on Mar. 1, 1995.
Ajemba et al., “Application of the Shrinking Core Model to the Analysis of Alumina Leaching From Ukpor Clay Using Nitric Acid”, International Journal of Engineering Research & Technology (IJERT), ISSN: 2278-0181, vol. 1 Issue 3, May 2012.
Zhou et al., “Extraction of Scandium from red mud by modified activated carbon and kinetics study”, Rare Metals, vol. 27, No. 3, Jun. 2008, pp. 223-227.
Ouellet, Dissertation 9689, (Extraction de l'alumine de l'argile de la region de Murdochville, Québec, Canada), “Extraction of Alumina from Clay in the Murdochville region of Quebec”, Canada—Universite Laval—Original French Version, Oct. 2004.
Ouellet, Dissertation 9689, (Extraction de l'alumine de l'argile de la region de Murdochville, Québec, Canada), “Extraction of Alumina from Clay in the Murdochville region of Quebec”, Canada—Universite Laval—English Translation, Oct. 2004.
Translator Certification—Sep. 10, 2012.
English Abstract BE1019347, “Hydrometallurgical Reactor”, published on Jun. 5, 2012.
Abstract Kao et al., “Solvent extraction of La(III) and Nd(III) from nitrate solutions with 2-ethylhexylphosphonic acid mono-2-ethylhexyl ester”, Chemical Engineering Journal, vol. 119, Issues 2-3, Jun. 15, 2006, pp. 167-174.
English Abstract CN102690954, “Back extraction and removement method for aluminium”, published on Sep. 26, 2012.
English Abstract CN101781719, “Method for recovering rare earth from oil shale waste slag”, published on Jul. 21, 2010.
English Abstract CN102643985, “Method for extracting valuable metals from high-iron bauxite with step-by-step acid leaching”, published on Aug. 22, 2012.
English Abstract CN102628105, “Method for comprehensively recycling and using baric waste slag in refined aluminum production process”, published on Aug. 8, 2012.
English Abstract of CN103964480(A), “Process for Producing Aluminum Oxide by Using Hydrochloric Acid Method”, published on Aug. 6, 2014.
English Abstract of CN203922759(U), “Device for Preparing Aluminum Oxide by Roasting Aluminum Chloride Hexahydrate Crystal through Rotary Kiln”, published on Nov. 5, 2014.
English Abstract of EP0850881(A1), “Process and Apparatus for the Preparation of Iron Oxides from Solutions Containing Hydrochloric Acid Iron Oxide Chloride”, published on Jul. 1, 1998.
English Abstract of JPH0543252A, “Method for Removing Halogen Radical in Ferric Oxide”, published on Feb. 23, 1993.
English Abstract of JPH0656429(A), “Production of Plate-Like Iron Oxide Particulate Powder”, published on Mar. 1, 1994.
English Abstract of WO2008070885(A2), “Method for Increasing the Specific Surface of Iron Oxides in Spray Roasting Plants”, published on Jun. 19, 2008.
Partial Translation of Abtract and Claims of CN1099424A, “Method for Treating Blast Furnace Slag with Diluted Chlorhydric Acid”, published on Mar. 1, 1995.
English Abstract—Chi et al., Derwent Acc-No. 2010-L68306 for the patent family including CN 101811712 A, published on Aug. 25, 2010.
Perander et al., “The Nature and Impacts of Fines in Smelter-Grade Alumina”, Journal of Minerals, Metals & Materials Society 61.11 (2009): 33-39. Springer Link. Web. Aug. 12, 2015. <http://link.springer.com/article/10.1007/02Fs11837-009-0164-x>.
English Abstract of JPH04354836(A), “Method for Leaching Silicomagnesionickel Ore”, Published on Dec. 9, 1992.
An English translation of Zhang et al., “Research on the Kinetics of Alumina from Kaolinite Leaching in Sulphuric Acid”, Journal of Hefei University of Technology, vol. 24, No. 1, Feb. 2001, pp. 71-74.
An English translation of Zhang et al., “Research of the Controlling Steps of the Reaction of Kaolin and Hydrochloric Acid”, Journal of Hefei University of Technology, vol. 21, No. 1, Feb. 1998, pp. 50-53.
An English translation of Zhang et al., “Kinetics Research on Alumina in Kaolinite Leached by Hydrochloric Acid”, Journal of Hefei University of Technology, vol. 22, No. 2, Apr. 1999, pp. 33-36.
Certification of translation from Park IP Translations dated May 14, 2012.
An English Abstract of CN101450811 “Method for extracting alumina from coal gangue”, published on Jun. 10, 2009.
An English Abstract of JP2001162108 “Method for Manufacturing Iron-Aluminum Combined Flocculant”, published on Jun. 19, 2001.
English Translation of Abstract of CN101462757, “Preparation of nano Na-beat-alumina powder”, Jun. 24, 2009.
An English Abstract of JP57145027 “Preparation of Granular Alumina”, published on Sep. 7, 1982.
An English Abstract of CA1065068 “Method of Selectively Precipitating Metals From Solutions”, published on Oct. 23, 1979.
English Translation of Abstract of CN101289705, “Process for abstracting vanadium from iron-smeltin waste slag of vanadium-containing iron ore”, Jul. 14, 2010.
English Translation of Abstract of CN102220487, “Method for extracting vanadium and aluminum from vanadium-containing stone coal and clay vanadium ore”, Oct. 19, 2011.
English Translation of Abstract of CN102241410, “Ecological and Comprehensive Utilization Method of Coal Ash”, Nov. 16, 2011.
English Translation of Abstract of RU2363748, “Method of Producing Aluminium”, Aug. 10, 2009.
English Translation of Abstract of ES2194586, “Separation procedure for contaminatory metals present in acid solutions involves liquid-liquid extraction with mixtures based on phosphonated dialkyl alkyl, trialkyl phosphates and acid phosphates”, Mar. 1, 2005.
English Translation of Abstract of WO2009005115, “Composition for promotion of reduction in size of adipocyte”, Jan. 8, 2009.
English Translation of Abstract of CN101434484, “Processes for producing alumina ceramic valve body and use thereof”, May 20, 2009.
English Translation of Abstract of CN101045543, “Method for preparing sheet alumina using coal series kaolin rock or flyash as raw material”, Oct. 3, 2007.
English Translation of Abstract of CN101249965, “Method for preparing ultra-fine white carbon black and nano alumina by using kaolinite as raw material”, Aug. 27, 2008.
English Translation of Abstract of CN101045538, “Method for preparing modified silicon oxide using coal series kaolin rock or flyash”, Oct. 3, 2007.
Tceisele, “Primary Metal Production”, Dec. 3, 2007.
US EPA, “Alumina & Aluminum”, Office of Resource Conservation and Recovery, Apr. 2, 2012.
Wahab et al., “Alumina Recovery From Iraqi Kaolinitic Clay by Hydrochloric Acid Route”, Iraqi Bulletin of Geology and Mining, vol. 2, No. 1, 2006, pp. 67-76.
Abstract of Dash et al., “Acid dissolution of alumina from waste aluminium dross”, Hydrometallurgy, vol. 92, issues 1-2, May 2008, pp. 48-53.
Copson et al., “Extraction of Alumina from Clays by the Lime-sinter Modification of the Pedersen Process”, New York Meeting, Feb. 1944, pp. 241-254.
Dutrizac et al., “The Precipitation of Hematite from Ferric Chloride Media at Atmospheric Pressure”, Mining and Mineral Sciences Laboratories, vol. 30B, Dec. 1999, pp. 993-1001.
English Abstract of JP10121164, published on May 12, 1998.
English Abstract of EP1817437, published on Aug. 15, 2007.
English Abstract of CN2292806, published on Sep. 30, 1998.
English Abstract of CN1986895, published on Jun. 27, 2007.
English Abstract of CN1796608, published on Jul. 5, 2006.
English Abstract of CN1699609, published on Nov. 23, 2005.
English Abstract of CN1410599, published on Apr. 16, 2003.
English Abstract of CN1397653, published on Feb. 19, 2003.
English Abstract of CN1192479, published on Sep. 9, 1998.
English Abstract of CN1478600, published on Mar. 3, 2004.
English Abstract of CN1140148, published on Jan. 15, 1997.
English Abstract of CN1127791, published on Jul. 31, 1996.
English Abstract of CN1131200, published on Sep. 18, 1996.
English Abstract of CN1061246, published on May 20, 1992.
English Abstract of CN1043752, published on Jul. 11, 1990.
English Abstract of CN102153128, published on Aug. 17, 2011.
English Abstract of CN102139943, published on Aug. 3, 2011.
English Abstract of CN102127641, published on Jul. 20, 2011.
English Abstract of CN102071317, published on May 25, 2011.
English Abstract of CN102071315, published on May 25, 2011.
English Abstract of CN102061392, published on May 18, 2011.
English Abstract of CN102030355, published on Apr. 27, 2011.
English Abstract of CN102021343, published on Apr. 20, 2011.
English Abstract of CN102011010, published on Apr. 13, 2011.
English Abstract of CN101824555, published on Sep. 8, 2010.
English Abstract of CN1045812, published on Oct. 3, 1990.
English Abstract of CN101407879, published on Apr. 15, 2009.
English Abstract of CN101307384, published on Nov. 19, 2008.
English Abstract of CN101161834, published on Apr. 16, 2008.
English Abstract of CN1844421, published on Oct. 11, 2006.
English Abstract of CN101182601, published on May 21, 2008.
English Abstract of CN1043752C, published on Jun. 23, 1999.
English Abstract of CN87101034, published on Aug. 28, 1991.
English Abstract of RU2079431, published on May 20, 1997.
English Abstract of RU2063458, published on Jul. 10, 1996.
English Abstract of RU 2 048 565, published on Nov. 20, 1995.
English Abstract of RU 2 040 587, published on Jul. 25, 1995.
English Abstract of RU 2 034 074, published on Apr. 30, 1995.
English Abstract of RU 2 031 168, published on Mar. 20, 1995.
English Abstract of RU 2 020 175, published on Sep. 30, 1994.
English Abstract of RU 2 010 876, published on Apr. 15, 1994.
English Abstract of RU 2 094 374, published on Oct. 27, 1997.
English Abstract of RU 2 081 831, published on Jun. 20, 1997.
English Abstract of RU 2 070 596, published on Dec. 20, 1996.
English Abstract of WO 2007122720, published on Nov. 1, 2007.
English Abstract of WO 2004085719, published on Oct. 7, 2004.
English Abstract of SU 1 734 395, published on Oct. 27, 1996.
English Abstract of RU2416655, published on Apr. 20, 2011.
English Abstract of RU2008113385, published on Oct. 20, 2009.
English Abstract of RU2361941, published on Jul. 20, 2009.
English Abstract of RU2257348, published on Jul. 27, 2005.
English Abstract of RU2247788, published on Mar. 10, 2005.
English Abstract of RU2236375, published on Sep. 20, 2004.
English Abstract of RU2205242, published on May 27, 2003.
English Abstract of RU2201988, published on Apr. 10, 2003.
English Abstract of RU2196184, published on Jan. 10, 2003.
English Abstract of RU2189358, published on Sep. 20, 2002.
English Abstract of RU2176680, published on Dec. 10, 2001.
English Abstract of RU2162898, published on Feb. 10, 2001.
English Abstract of RU2162112, published on Jan. 20, 2001.
English Abstract of RU2158170, published on Oct. 27, 2000.
English Abstract of RU2147623, published on Apr. 20, 2000.
English Abstract of RU2147622, published on Apr. 20, 2000.
English Abstract of RU2140998, published on Nov. 10, 1999.
English Abstract of RU2119816, published on Oct. 10, 1998.
English Abstract of KR20070028987, published on Mar. 13, 2007.
English Abstract of JP9324227, published on Dec. 16, 1997.
English Abstract of JP9324192, published on Dec. 16, 1997.
English Abstract of JP9291320, published on Nov. 11, 1997.
English Abstract of JP9249672, published on Sep. 22, 1997.
English Abstract of JP9248463, published on Sep. 22, 1997.
English Abstract of JP9208222, published on Aug. 12, 1997.
English Abstract of JP9194211, published on Jul. 29, 1997.
English Abstract of JP9176756, published on Jul. 8, 1997.
English Abstract of JP9143589, published on Jun. 3, 1997.
English Abstract of JP8232026, published on Sep. 10, 1996.
English Abstract of JP5051208, published on Mar. 2, 1993.
English Abstract of JP4198017, published on Jul. 17, 1992.
English Abstract of JP4183832, published on Jun. 30, 1992.
English Abstract of JP4046660, published on Feb. 17, 1992.
English Abstract of JP3173725, published on Jul. 29, 1991.
English Abstract of JP2179835, published on Jul. 12, 1990.
English Abstract of JP2080530, published on Mar. 20, 1990.
English Abstract of JP2011116622, published on Jun. 16, 2011.
English Abstract of JP2011046588, published on Mar. 10, 2011.
English Abstract of JP2010270359, published on Dec. 2, 2010.
English Abstract of JP2008194684, published on Aug. 28, 2008.
English Abstract of JP2007327126, published on Dec. 20, 2007.
English Abstract of JP2007254822, published on Oct. 4, 2007.
English Abstract of JP2006348359, published on Dec. 28, 2006.
English Abstract of JP2006028187, published on Feb. 2, 2006.
English Abstract of JP2005139047, published on Jun. 2, 2005.
English Abstract of JP2000313928, published on Nov. 14, 2000.
English Abstract of JP10158629, published on Jun. 16, 1998.
English Abstract of AU2008286599A1, “A process of smelting monazite rare earth ore rich in Fe”, published on Feb. 19, 2009.
English Abstract of KR820001546, “Production of Titanium Metal Valves”, published on Aug. 31, 1982.
English Abstract of KR100927466, published on Nov. 19, 2009.
English Abstract of RU 2 069 180, published on Nov. 20, 1996.
English Abstract of RU 2 068 392, published on Oct. 27, 1996.
English Abstract of RU 2 062 810, published on Jun. 27, 1996.
English Abstract of RU 2 055 828, published on Mar. 10, 1996.
English Abstract of RU 2 049 728, published on Dec. 10, 1995.
English Abstract of RU 1 704 483, published on Oct. 27, 1996.
English Abstract of SU1567518, published on May 30, 1990.
English Abstract of SU1424174, published on Jul. 23, 1991.
English Abstract of RU2038309, published on Jun. 27, 1995.
English Abstract of HU51574, published on May 28, 1990.
English Abstract of JP2008253142, published on Oct. 23, 2008.
English Abstract of JP2005152756, published on Jun. 16, 2005.
English Abstract of JP2005082462, published on Mar. 31, 2005.
English Abstract of JP2005219938, published on Aug. 18, 2005.
English Abstract of JP2004036003, published on Feb. 5, 2004.
English Abstract of WO2011092292, published on Aug. 4, 2011.
English Abstract of CN101157453, published on Apr. 9, 2008.
English Abstract of AU2737892, published on May 13, 1993.
English Abstract of JPH09249420, published on Sep. 22, 1997.
English Abstract of CN102515590, published on Jun. 27, 2012.
English Abstract of CN102502745, published on Jun. 20, 2012.
English Abstract of WO2007082447, published on Jul. 26, 2007.
English Abstract of CN102694218, published on Sep. 26, 2012.
English Abstract of RU2183225, published on Jun. 10, 2002.
English Abstract of JP2009249674, published on Oct. 29, 2009.
English Abstract of CN102719674, published on Oct. 10, 2012.
English Abstract of 102680423, published on Sep. 19, 2012.
English Abstract of JP2005112636, published on Apr. 28, 2005.
English Abstract of CN101285127, published on Oct. 15, 2008.
Australie Minerals & Mining Group LTD, “AMMG Updates Process Design for HPA Chemical Project”, Oct. 3, 2014, pp. 1-4.
English Abstract of CN1044126(C), “Stretched polypropylene film”, published on Jul. 14, 1999.
English Abstact of RU2237111 (C1), “Method of Recovering Magnesium From Silicon-Containing Wastes”, published on Sep. 27, 2004.
English Translation of CN102452677, published on May 16, 2012.
English Abstract of CN101781719, “Method for recovering rare earth from oil shale waste slag”, published on Jul. 21, 2010.
English Abstract of RU2158787, “Process of Winning of Magnesium”, published on Nov. 10, 2000.
Related Publications (1)
Number Date Country
20150225808 A1 Aug 2015 US
Provisional Applications (4)
Number Date Country
61705898 Sep 2012 US
61713795 Oct 2012 US
61726971 Nov 2012 US
61837715 Jun 2013 US