Recovery of metals from spent supported catalysts

Abstract

An improved method for recovering metals from spent supported catalysts, including spent supported hydroprocessing catalysts. The method, and related processes including the method, are useful for recovering spent supported catalyst metals used in the petroleum and chemical processing industries. The method generally involves a combination of pyrometallurgical and hydrometallurgical processes, and includes forming potassium carbonate calcine from spent supported catalyst containing Group VIIIB / VIB / VB metal compound(s) combined with potassium carbonate, and extracting and recovering soluble Group VIB and soluble Group VB metal compounds from the potassium carbonate calcine.

Description

[Technical field]

[0001] CROSS-REFERENCE TO RELATED APPLICATIONS This application claims the benefit of priority to U.S. Provisional Application No. 63 / 310,034, filed February 14, 2022, entitled "METALS RECOVERY FROM SPENT SUPPORTED CATALYST," and U.S. Patent Application No. 17 / 744,018, filed May 13, 2022, the disclosures of which are incorporated by reference in their entireties herein.

[0002] The present invention relates to a process for recovering metals from spent supported catalysts, including spent supported hydroprocessing catalysts. [Background technology]

[0003] Catalysts have been widely used in the petroleum refining and chemical processing industries for many years. Hydroprocessing catalysts, including hydrotreating catalysts and hydrocracking catalysts, are now widely used in facilities around the world. Spent hydroprocessing catalysts, or "spent" hydroprocessing catalysts, that are no longer sufficiently active (or need to be replaced for another reason), typically contain metal components such as molybdenum, nickel, cobalt, vanadium, and the like. Many such catalysts include support materials, and "supported" catalysts in the broad sense. In contrast, catalysts without such support materials are typically classified as unsupported or bulk catalysts, or may simply not refer to the use of a support material.

[0004] The advent of heavier feedstocks has forced refineries to use more catalysts in hydrotreating to remove sulfur and contaminants from the feedstocks. These catalytic processes generate significant amounts of spent catalysts, which serve two purposes: they contain valuable metals and, in keeping with environmental concerns, they keep them from being landfilled.

[0005] Various processes for recovering catalytic metals from spent catalysts have been described in the literature. For example, US Patent Application Publication No. 2007 / 0025899 discloses a process for recovering metals such as molybdenum, nickel and vanadium from spent catalysts with multiple steps and equipment for recovering metal complexes of molybdenum and nickel. US Patent No. 6,180,072 discloses another complex process requiring an oxidation step and solvent extraction to recover metals from spent catalysts containing at least metal sulfides. US Patent No. 7,846,404 discloses a process using pH adjustment and precipitation to recover metals from ammonia pressure leach solutions produced during oxidative pressure leaching of spent catalysts.

[0006] More recent attempts are disclosed, for example, in WO2021 / 005526A1, in which the use of sodium carbonate is generally described in the recovery of metals from spent slurry (unsupported) catalysts. WO2021 / 150552A1 further generally describes the use of potassium carbonate in the recovery of metals from spent slurry (unsupported) catalysts. Additional publications of interest are set out in Appendix A - Publications.

[0007] Despite advances made in the recovery of catalytic metals from spent catalysts, particularly from spent unsupported catalysts, there continues to be a need for improved and simplified processes for recovering catalytic metals (including, but not limited to, molybdenum, and vanadium) from spent catalysts, particularly from spent supported catalysts. Summary of the Invention

[0008] The present invention is directed to a method for recovering catalytic metals from spent catalysts, particularly spent supported hydroprocessing catalysts. One of the goals of the present invention is to provide an improved spent supported catalyst metals recovery process that reduces capital and operating costs in metal recovery, preferably with improved metal recovery efficiency. The present invention provides an innovative and cost-effective approach in the recovery of supported catalytic metals, while also providing an improvement in supported catalyst metals recovery in general, which addresses important environmental sustainability needs in the oil and gas industry, as well as the metals recovery industry.

[0009] An improved method for recovering metals from spent supported catalyst is disclosed. The method and related processes including the method are useful for recovering supported catalyst metals used in the petroleum and chemical processing industries. The method generally includes both pyrometallurgical and hydrometallurgical techniques and methods. The pyrometallurgical method includes oxidative roasting of a mixture of degreased spent supported catalyst combined with potassium carbonate to form the calcine of the mixture. The calcine is then leached with water to provide soluble Group VB and Group VIB metal compounds and a residue including Group VB, Group VIB and Group VIIIB metal compounds. The soluble fraction including Group VB and Group VIB metal compounds is combined with ammonium salts to convert the Group VB and Group VIB metals to their ammonium forms under metathesis reaction conditions. The Group VB metal is recovered by subsequent crystallization and the remaining Group VIB metal is recovered by acidification and precipitation of the Group VIB metal.

[0010] In one aspect, the pyrometallurgical process includes heating a mixture of deoiled, spent supported catalyst comprising Group VIB, Group VIIIB, and Group VB metals in combination with potassium carbonate under oxidizing conditions to reduce the levels of sulfur and carbon present in the catalyst and form a spent supported catalyst / potassium carbonate calcine comprising water soluble Group VIB metal compounds, water soluble Group VB metal compounds, and water insoluble Group VIIIB metal compounds; and combining the spent supported catalyst / potassium carbonate calcine with water under slurry leach process conditions to form a spent supported catalyst / potassium carbonate calcine slurry. leaching the water-soluble Group VIB metal compounds and the water-soluble Group VB metal compounds from the spent supported catalyst / potassium carbonate calcine slurry; separating and removing a filtrate and a solid residue from the spent supported catalyst / potassium carbonate calcine slurry, the filtrate comprising the water-soluble Group VIB metal compounds and the water-soluble Group VB metal compounds and the solid residue comprising the water-insoluble Group VIIIB metal compounds; and recovering the water-soluble Group VIB metal compounds and the water-soluble Group VB metal compounds from the filtrate of the spent supported catalyst / potassium carbonate calcine slurry.

[0011] In a further aspect, the hydrometallurgical process comprises separately recovering Group VIB and Group VB metal compounds obtained from the spent supported catalyst from a solution containing Group VIB and Group VB metal compounds by contacting a mixture of the Group VIB and Group VB metal compounds with an ammonium salt under metathesis reaction conditions effective to convert the Group VIB and Group VB metal compounds to ammonium Group VB and ammonium Group VB metal compounds, subjecting the mixture containing the ammonium Group VB metal compound to conditions effective to crystallize the ammonium Group VB metal compound, and washing the crystallized ammonium Group VB metal compound with a saturated ammonium Group VB metal compound wash solution. heating the ammonium Group VB metal compound under conditions effective to liberate ammonia and separately recovering the Group VB metal compound and the ammonia; contacting the ammonium Group VIB metal compound filtrate with an inorganic acid under conditions effective to form a Group VIB metal oxide compound precipitate, and an ammonium salt of the inorganic acid; filtering and washing the Group VIB metal oxide compound precipitate with an ammonium Group VIB metal oxide compound wash solution at a predetermined wash temperature and recovering the Group VIB metal oxide compound precipitate.

[0012] It should be understood that the scope of the present invention is not limited by any of the representative drawings attached to this disclosure, but is defined by the claims of this application. [Brief description of the drawings]

[0013] [Figure 1] 1 is a general block diagram of an embodiment of a roasting and extraction process for recovering metals from deoiled spent supported catalyst in accordance with the present invention.

[0014] [Diagram 2] 1 is a general block diagram of an embodiment of a water recovery process for recovering metals from deoiled spent supported catalyst in accordance with the present invention.

[0015] [Diagram 3] FIG. 1 is a general block diagram of an embodiment of a combined roasting / extraction and water recovery process for recovering metals from deoiled spent supported catalyst in accordance with the present invention.

[0016] [Figure 4] 13 provides heat and off-gas results from examples. [Diagram 5] 13 provides heat and off-gas results from examples. [Figure 6] 13 provides heat and off-gas results from examples. [Figure 7] 13 provides heat and off-gas results from examples. [Figure 8] 13 provides heat and off-gas results from examples. [Figure 9] 13 provides heat and off-gas results from examples. DETAILED DESCRIPTION OF THE PREFERRED EMBODIMENTS

[0017] Although exemplary embodiments of one or more aspects are provided herein, the disclosed processes may be implemented using any number of methods. The present disclosure is not limited to the exemplary or specific embodiments, drawings, and techniques illustrated herein, including any exemplary designs and embodiments illustrated and described herein, but may be modified within the scope of the appended claims, along with the full range of equivalents.

[0018] The following terms, nomenclature, and definitions are applicable to this disclosure unless otherwise specified. If a term is used in this disclosure but not specifically defined herein, the definition from the IUPAC Compendium of Chemical Terminology, 2nd ed (1997) may be applied, provided that such definition does not conflict with any other disclosure or definition applied herein or render any claim to which it applies unclear or unacceptable. In the event that any definition or usage set forth in any document incorporated herein by reference conflicts with the definition or usage set forth herein, the definition or usage set forth herein shall apply.

[0019] "Supported catalyst" generally refers to a catalyst composition having a "support" material, including conventional catalyst forms, including preformed shaped catalyst supports, on which metals are then loaded by impregnation or deposition to form a supported catalyst. The term "support", particularly when used in the terms "supported catalyst" and "catalyst support", refers to conventional materials, usually solids with a large surface area, that support the catalytic material. The support material may be inert or may participate in the catalytic reaction, and may be porous or non-porous. Typical catalyst supports include various types of carbon, alumina, silica and silica-alumina, such as amorphous silica aluminates, zeolites, alumina-boria, silica-alumina-magnesia, silica-alumina-titania, and materials obtained by adding other zeolites and other composite oxides.

[0020] "Slurry catalyst" may be used interchangeably with "bulk catalyst", "unsupported catalyst" or "self-supported catalyst" to mean that the catalyst composition is not a traditional catalyst form having a preformed shaped catalyst support that is later loaded with metals by impregnation, i.e., a deposited catalyst. Such bulk catalysts may be formed by precipitation or may have a binder incorporated into the catalyst composition. Slurry or bulk catalysts may be formed from metal compounds without any binder. In the slurry form, such catalysts include dispersed particles in a liquid mixture such as a hydrocarbon oil, i.e., a "slurry catalyst."

[0021] "Heavy oil" feedstocks or feedstocks refer to heavy and extra heavy crude oils, including but not limited to residual oils, coal, bitumen, tar sands, oils obtained from pyrolysis of waste, polymers, biomass, coke and oil shale. Heavy oil feedstocks can be liquid, semi-solid and / or solid. Examples of heavy oil feedstocks include but are not limited to Canadian tar sands, Santos and Campos Basins of Brazil, Gulf of Suez of Egypt, Chad, Zulia of Venezuela, Malaysia and Sumatra vacuum residua of Indonesia. Other examples of heavy oil feedstocks include "bottoms of the barrel" and "residues" (i.e., "residual oils"), residues left over from refining processes, including atmospheric bottoms having a boiling point of at least 650°F (343°C), vacuum bottoms having a boiling point of at least 975°F (524°C), or "resid pitches" and "vacuum residues" having boiling points above 975°F (524°C).

[0022] "Processing," "processed," "upgrading," "upgrading," and "upgraded," when used in conjunction with heavy oil feedstocks, describe a heavy oil feedstock that has been hydrotreated or has been hydrotreated, or a resulting material or crude product, in which the molecular weight of the heavy oil feedstock has been reduced, the boiling range of the heavy oil feedstock has been reduced, the concentration of asphaltenes has been reduced, the concentration of hydrocarbon free radicals has been reduced, and / or the amount of impurities such as sulfur, nitrogen, oxygen, halides, and metals has been reduced.

[0023] The upgrading or processing of heavy oil feedstocks is generally referred to herein as "hydroprocessing" (hydrocracking or hydroconversion). Hydroprocessing is meant as any process carried out in the presence of hydrogen, including, but not limited to, hydroconversion, hydrocracking, hydrogenation, hydrotreating, hydrodesulfurization, hydrodenitrogenation, hydrodemetallization, hydrodearomatization, hydroisomerization, hydrodewaxing, and hydrocracking (including selective hydrocracking).

[0024] The terms "Hydrogen" or "hydrogen" refer to hydrogen itself and / or the compound(s) that provide a source of hydrogen.

[0025] "Hydrocarbon-containing," "hydrocarbon," and similar terms refer to compounds containing only carbon and hydrogen atoms. When a particular group is present in the hydrocarbon, other identifiers may be used to indicate the presence of that particular group (e.g., halogenated hydrocarbon indicates the presence of one or more halogen atoms replacing an equal number of hydrogen atoms in the hydrocarbon).

[0026] "Spent supported catalyst" refers to a supported catalyst that has been used in a hydroprocessing operation and has lost its activity. In general, a catalyst may be referred to as a "spent" catalyst if the reaction rate constant of the catalyst is less than a certain designated value at a designated temperature compared to a fresh catalyst. In some circumstances, a catalyst may be a "spent" catalyst if its reaction rate constant is 80% or less, or in another embodiment, perhaps 50% or less, relative to a fresh unused catalyst. In one embodiment, the metal component of the spent catalyst includes at least one of Group VB, Group VIB, and Group VIIIB metals (of the periodic table), such as vanadium (V), molybdenum (Mo), tungsten (W), iron (Fe), nickel (Ni), and cobalt (Co). The most commonly found metal to be recovered is Mo. Although not necessarily the case, the spent supported catalyst typically includes sulfides of Mo, Ni, and V.

[0027] "Deoiled spent supported catalyst" generally refers to a "spent supported catalyst" as described above that has been subjected to a deoiling process. Generally, deoiled spent supported catalyst contains some residual hydrocarbons, such as unconverted oil and / or hydroprocessing products, as well as other chemical compounds and materials. For example, deoiled spent supported catalyst may typically contain 15 wt% or more of residual hydrocarbons, or may have reduced amounts (such as 1 wt% or less or 1000 ppm or less) if treated to remove such hydrocarbons. The specified content of such additional components, whether in general or specific terms, is specified herein as appropriate.

[0028] "Metal" refers to a metal in its elemental, compound or ionic form. "Metal precursor" refers to a metal compound that is subjected to a method or process. The terms "metal", "metal precursor" or "metal compound" in the singular form are not limited to a single metal, metal precursor or metal compound, e.g., Group VIB, VIII or V metals, but also include plural references in mixtures of metals. The terms "soluble" and "insoluble", with respect to Group VIB, VIII or V metals or metal compounds, mean that the metal component is in protic liquid form or that the metal or metal compound is soluble or insoluble in the specified process or solvent, unless otherwise stated.

[0029] "Group IIB" or "Group IIB metals" refers to zinc (Zn), cadmium (Cd), mercury (Hg), and combinations thereof, in either elemental, compound, or ionic form.

[0030] "Group IVA" or "Group IVA metal" refers to germanium (Ge), tin (Sn), or lead (Pb), and combinations thereof, in either elemental, compound, or ionic form.

[0031] "Group VB metal" refers to vanadium (V), niobium (Nb), tantalum (Ta), and combinations thereof, in elemental, compound, or ionic form.

[0032] "Group VIB" or "Group VIB metals" refers to chromium (Cr), molybdenum (Mo), tungsten (W), and combinations thereof, in either elemental, compound, or ionic form.

[0033] "Group VIIIB" or "Group VIIIB metals" refers to iron (Fe), cobalt (Co), nickel (Ni), ruthenium (Ru), rhenium (Rh), rhodium (Rh), palladium (Pd), osmium (Os), iridium (Ir), platinum (Pt), and combinations thereof, in either elemental, compound, or ionic form.

[0034] References to Mo or "molybdenum" are merely for illustrative purposes as a Group VIB metal and are not intended to exclude other Group VIB metals / compounds and mixtures of Group VIB metals / compounds. Similarly, references to Ni or "nickel" are merely for illustrative purposes and are not intended to exclude other Group VIIIB non-noble metal components, Group VIIIB metals, Group VIB metals, Group IVB metals, Group IIB metals and mixtures thereof that may be used in hydroprocessing catalysts. Similarly, references to V or "vanadium" are merely for illustrative purposes as to any Group VB metal components that may be present in the spent supported catalyst and are not intended to exclude other Group VB metals / compounds and mixtures that may be present in the spent supported catalyst used for metal recovery.

[0035] The description of combinations of metal compounds, represented by the use of the term "Group VIIIB / Group VIB / Group VB" to describe metal compounds that may be present, is intended to mean that Group VIIIB metal compounds, Group VIB metal compounds, or Group VB metal compounds, and any combination thereof, may be present. For example, if the spent catalyst contains metal compounds of Mo, V, Ni, and Fe as oxygen and / or sulfur-containing compounds, the term "Group VIIIB / Group VIB / Group VB" should be understood to include single metal compounds and mixed metal compounds, i.e., metal compounds containing Group VIIIB metals, Group VIB metals, Group VB metals, or combinations thereof. Representative compounds include, for example, MoS2, V2S3, NiS, FeS, MoO3, VO3, NiO, VO5, Fe2O3, NiMoO4, FeVO4, Ni3(VO4)2, and the like. Similarly, the terms "Group VB / Group VIB" metal(s) and metal oxide(s) refer to metal or metal oxide compounds that include a Group VB metal, a Group VIB metal, or a combination thereof.

[0036] "Molecular sieve" refers to a material having uniform pores of molecular dimensions within a framework structure; depending on the type of molecular sieve, only certain molecules can access the pore structure of the molecular sieve, while other molecules are excluded, for example, due to molecular size and / or reactivity. Zeolites, crystalline aluminophosphates, and crystalline silicoaluminophosphates are representative examples of molecular sieves.

[0037] Although compositions and methods or processes are often described in this disclosure in terms of "comprising" various components or steps, the compositions and methods may "consist essentially of" or "consist of" the various components or steps, unless otherwise specified.

[0038] The terms "a," "an," and "the" are intended to include multiple alternatives, e.g., at least one. For example, disclosure of a "transition metal" or an "alkali metal" is meant to encompass one transition metal or alkali metal, or mixtures or combinations of multiple transition metals or alkali metals, unless otherwise specified.

[0039] Any numerical values ​​within the detailed description and claims herein are modified by the term "about" or "approximately" the stated value to account for experimental error and variations that would be expected by one of ordinary skill in the art.

[0040] The present invention is a method for recovering metals from deoiled spent catalyst, the spent catalyst comprising Group VIB, Group VIIIB and Group VB metals. In one aspect, the method comprises: combining a deoiled spent supported catalyst containing a Group VIIIB metal, a Group VIB metal, and a Group VB metal with potassium carbonate to form a deoiled spent catalyst / potassium carbonate mixture; heating the deoiled spent supported catalyst / potassium carbonate mixture under oxidizing conditions to reduce the levels of sulfur and carbon and form a spent supported catalyst / potassium carbonate calcine containing water soluble Group VIB metal compounds, water soluble Group VB metal compounds, and water insoluble Group VIIIB metal compounds; combining the spent supported catalyst / potassium carbonate calcine with water under slurry leaching process conditions to form a spent supported catalyst / potassium carbonate calcine slurry and leaching water soluble Group VIB metal compounds and water soluble Group VB metal compounds from the spent supported catalyst / potassium carbonate calcine; Separating and removing a filtrate and a solid residue from the spent supported catalyst / potassium carbonate calcine slurry, the filtrate comprising water-soluble Group VIB metal compounds and water-soluble Group VB metal compounds, and the solid residue comprising water-insoluble Group VIIIB metal compounds;

[0041] and recovering the water-soluble Group VIB metal compound and the water-soluble Group VB metal compound from the filtrate obtained from the spent supported catalyst / potassium carbonate calcine slurry.

[0042] The method provides improved spent supported catalyst metal recovery and a simplified, cost-effective approach in the recovery of metals from spent supported catalysts. The method utilizes a leaching extraction step, i.e., aqueous leaching extraction of the spent supported catalyst / potassium carbonate calcine formed from heating the deoiled spent supported catalyst / potassium carbonate mixture under oxidizing conditions. The method uses a single-step leaching extraction, which includes aqueous leaching extraction of potassium carbonate calcine formed from combining the calcined spent supported catalyst with potassium carbonate. The method does not require the use of additional extraction steps (within the method), such as adding other solvents (e.g., tertiary amines for extraction of soluble metals), or additional treatment of organic compounds (e.g., with activated carbon, bentonite) in combination with the use of potassium carbonate.

[0043] The spent supported catalyst generally originates from a supported Group VIB metal sulfide catalyst optionally containing metals selected from Group VB metals such as V, Nb, Group VIIIB metals such as Ni, Co, Group VIIIB metals such as Fe, Group IVB metals such as Ti, Group IIB metals such as Zn, and combinations thereof. Certain additional metals may be added to the supported catalyst formulation to improve certain properties or to modify the activity and / or selectivity of the catalyst. Suitable supported catalysts include catalysts useful in various reactor types, for example, fixed bed or ebullated bed reactors.

[0044] Suitable catalysts for use as the spent supported catalyst in this process are disclosed in U.S. Pat. Nos. 11,001,502, 10,913,054, 10,213,772, 9,920,260, 9,206,361, 7,803,266, 7,185,870, 7,449,103, 8,024,232, 7,618,530, 6,589,908, and the like. Nos. 6,667,271, 7,642,212, 7,560,407, 6,030,915, U.S. Pat. No. 5,980,730, U.S. Pat. No. 5,968,348, U.S. Pat. No. 5,498,586, and U.S. Patent Publication Nos. 2011 / 0226667, 2009 / 0310435, and 2011 / 0306490.

[0045] Prior to metal recovery, the spent supported catalyst may be treated to remove residual hydrocarbons such as oil, precipitated asphaltenes, and other oil residues. The spent supported catalyst (before de-oiling) typically contains carbon fines, metal fines, and spent supported catalyst in unconverted residual hydrocarbon oil, with a solid content ranging from 5 to 50 wt%. The de-oiling process may include the use of a solvent to remove the oil, followed by liquid / solid separation to recover the de-oiled spent catalyst. The treatment process may further include a thermal treatment step, e.g., drying and / or pyrolysis, to remove the hydrocarbons from the spent supported catalyst. In another embodiment, the de-oiling may include the use of subcritical high pressure gas, optionally with surfactants and additives, to expel / remove the oil from the spent supported catalyst.

[0046] The deoiled spent supported catalyst typically contains less than 5 wt% of hydrocarbons as unconverted residuals, or more specifically, less than 2 wt% of hydrocarbons, or less than 1 wt% of hydrocarbons. In some cases, the deoiled spent supported catalyst is substantially free of residual hydrocarbons, free of residual hydrocarbons, or contains less than about 1000 ppm, less than 500 ppm, or less than 100 ppm of residual hydrocarbons. The amount of metals recovered from the deoiled spent supported catalyst generally depends on the composition of the catalyst used in hydroprocessing, such as a Group VIB metal sulfide catalyst, a bimetallic catalyst containing a Group VIB metal and a Group VIIIB metal, or a multimetallic catalyst containing at least Group VIB and other (e.g., promoter) metal(s). After the oil removal treatment process, the spent supported catalyst containing metals for recovery may be in the form of a coke-like material, which can be crushed accordingly, if necessary, for subsequent metal recovery processes.

[0047] Deoiling or removal of hydrocarbons from spent catalysts has been disclosed in a number of publications, including US7790646, US7737068, WO20060117101, WO2010142397, US20090159505A1, US20100167912A1, US20100167910A1, US20100163499A1, US20100163459A1, US20090163347A1, US20090163348A1 and US20060135631A1.

[0048] An illustrative example of a roasting and extraction method or process according to an embodiment of the invention is shown generally in Figure 1. A deoiled spent supported catalyst (11), e.g., a catalyst free or substantially free of residual hydrocarbons as described herein, is mixed (10) with potassium carbonate (12) and the mixture (13) is fed to a heating or roasting step (20) to reduce the sulfur and / or carbon content present in the supported catalyst to less than a predetermined amount, and then (23) forms a spent supported catalyst / potassium carbonate calcine from the heating / roasting step, calcination step (20). The heating / roasting may be performed as a separate batch or continuous process step in the same or different equipment. As noted above, the thermal oxidation products of sulfur and carbon from the catalyst may be utilized to establish the amount of time required for calcination (or completion of the calcination step). The spent supported catalyst calcine is then mixed (30) with water (31) to form a spent supported catalyst / potassium carbonate calcine slurry 33, which leaches the water-soluble Group VIB and Group VB metal compounds from the spent supported catalyst / potassium carbonate calcine. The spent supported catalyst / potassium carbonate calcine slurry (33) is separated (40) into a filtrate (43) and a residue (44). A deionized water wash (42) of about pH 10 of the residue (44) may be applied. The filtrate contains the soluble Group VIB and Group VB metals, and the Group VB and Group VB metal compounds are separated for subsequent recovery, e.g., as V2O5 and MoO3 in the case of vanadium and molybdenum. The residue (44), which contains insoluble compounds (e.g., Ni, Fe, residual Mo, V, and other metal compounds, etc.), may also be further processed for possible metal recovery or sent to a sugar refinery.

[0049] The heating / roasting stage (20 in FIG. 1) includes a first heating step to remove residual hydrocarbons, followed by a roasting step to form the calcine of the spent supported catalyst / potassium carbonate mixture. The heating step is carried out under inert conditions, e.g., at 500° C. for 1-2 hours. However, for deoiled spent supported catalysts having a low content of residual hydrocarbons, e.g., less than about 1000 ppm, such as may be the case with pre-treated supported catalysts, an initial heating step under inert conditions to remove residual hydrocarbons may not be necessary. Without being limiting, the heating may include, for example, slowly increasing the temperature to an initial temperature, e.g., a temperature in the range of 350-500° C., under an inert gas such as argon, for a period of time (e.g., 1-2 hours) suitable to remove residual hydrocarbons from pyrolysis. Following the heating step, a subsequent roasting step is carried out under oxidizing conditions to remove carbon and sulfur and form the spent catalyst / potassium carbonate calcine. Although shown as one step in FIG. 1, the heating and roasting steps may be carried out as separate, consecutive steps.

[0050] Calcination of the potassium carbonate / spent supported catalyst mixture, in step (20), is typically carried out under oxidizing conditions by increasing the temperature to a suitable calcination temperature, for example, in the range of 600°C to 650°C, or about 610°C to 650°C, or about 610°C to 630°C, or greater than about 600°C, or about 610°C, or about 620°C, or about 630°C, or about 640°C, or about 650°C, for a period of time suitable to form a calcined spent supported catalyst (e.g., typically more than 1 to 2 hours and less than about 24 hours, or more specifically, 0.5 to 12 hours, or 1 to 8 hours, or 4 to 8 hours, or about 12 hours, or 10 hours, or less than 8 hours). The heating step may also include first heating under inert gas heating conditions at a temperature of about 600°C or 550°C, or less than 500°C or 450°C for a predetermined time ranging from about 0.5 to 4 hours, or 1 to 3 hours, or less than about 4 hours, or less than 2 hours. In general, the degree of oxidation of the calcined spent supported catalyst may also be monitored by analysis of the thermal oxidation products of CO2 and SO2 during calcination to establish a suitable end point of the calcination. For example, the end point may be associated with a CO2 and SO2 level of less than about 1 wt%, less than about 0.8 wt%, less than about 0.5 wt%, less than about 0.2 wt%, or less than about 0.1 wt%. The thermal oxidation product gas may also be monitored by infrared detection for at least about 20 wt% O2, less than about 0.25 wt% CO2, and less than about 0.25 wt% SO2.

[0051] During the spent supported catalyst / potassium carbonate mixture calcination step (20 in FIG. 1), oxidative heating conditions may generally include heating in the presence of air, or a gas mixture containing sufficient oxygen and air to maintain oxidative conditions. If desired, modified oxidative conditions may be used, for example, an initial gas environment containing about 20 vol.% or less oxygen followed by gas conditions containing up to 80 vol.% oxygen.

[0052] Upon calcination of a spent supported catalyst / potassium carbonate mixture, for example when the supported catalyst contains Mo, Ni, V, Fe, C, and S, the following representative reactions are possible to form the following compounds and thermal oxidation product gases: MoS2+3K2CO3+9 / 2O2=K2MoO4+2K2SO4+3CO2 ↑ (1) V2S3+4K2CO3+7O2=2KVO3+3K2SO4+4CO2 ↑ (2) NiS+K2CO3+2O2=NiO+K2SO4+CO2 ↑ (3) 2FeS+2K2CO3+9 / 2O2=Fe2O3+2K2SO4+2CO2 ↑ (4) C+O2=CO2 ↑ (5) S+O2=SO2 ↑ (6) K2CO3+SO2+1 / 2O2=K2SO4+CO2 ↑ (7) NiMoO4+K2CO3=K2MoO4+NiO+CO2 ↑ (8) Ni3(VO4)2+K2CO3=2KVO3+3NiO+CO2 ↑ (9)

[0053] For example, if the supported catalyst contains Mo, Ni, V, Fe, C, and S, the spent catalyst combustion reaction in the presence of air may also form the following compounds and thermal oxidation product gases: V2S3+11 / 2O2=V2O5+3SO2 ↑ (10) MoS2+7 / 2O2=MoO3+2SO2 ↑ (11) 2FeS+7 / 2O2=Fe2O3+2SO2 ↑ (12) NiS+3 / 2O2=NiO+SO2 ↑ (13) C+O2=CO2 ↑ (14) S+O2=SO2 ↑ (15) V2S3+3NiS+10O2=Ni3(VO4)2+6SO2 ↑ (16) MoO3 + NiO = NiMoO4 (17)

[0054] Under some conditions, e.g., temperatures between 500-600°C, certain reactions involving nickel and vanadium may occur preferentially, leading to the formation of nickel vanadate. For example, reaction (16) may be dominant over reaction (13). The formation of spinel phases, e.g., NiMoO4, may also occur (reaction 17). The spinel phase is unavoidable, as its formation begins at about 250°C or temperatures before complete S and C oxidation can be fully realized.

[0055] Following calcination, the spent supported catalyst / potassium carbonate calcine is contacted with water to form a potassium carbonate calcine slurry. Leaching in water is performed to extract the soluble metal compounds to form a filtrate and an insoluble metal compound(s) residue including insoluble Group VIIIB / VIB / VB metal compound(s). The filtrate typically includes soluble molybdate and / or vanadate compounds, while the insoluble compounds typically include specific metal and mixed metal compounds. In the case of the representative reactions (1)-(9) above, such insoluble metal compounds are believed to include nickel oxide, nickel molybdate, and nickel vanadate mixed metal compounds.

[0056] The reaction of certain insoluble Group VB and Group VIB metal compounds (referred to as "spinel") with potassium carbonate may involve reactions (8) and (9) in the case of Mo and V metal compounds. NiMoO4+K2CO3→K2MoO4+NiO+CO2 ↑ (8) Ni3(VO4)2+K2CO3=2KVO3+3NiO+CO2 ↑ (9) For example, the refractory and non-leachable NiMoO4 formed from MoO3 and NiO (reaction 17) can both be at least partially reformed in the presence of alkali to water-soluble molybdates (reaction 8). Similarly, the refractory and non-leachable Ni3(VO4)2 formed from V and Ni sulfides (reaction 16) can both be at least partially reformed in the presence of alkali to water-soluble vanadates (reaction 9).

[0057] Typical leaching conditions include, but are not limited to, leaching temperatures in the range of about 60-90° C., 60-80° C., or 70-80° C., or greater than about 60° C. or greater than 70° C., leaching times in the range of about 1-5 hours, about 2-5 hours, or about 2-4 hours, and leaching pH in the range of about 9.5-11, about 10-11, or about 10-10.5. The spent supported catalyst / potassium carbonate calcine slurry leaching can be carried out with or without pH modification, so long as the pH of the calcine slurry is greater than 9.5.

[0058] Generally, extraction of up to about 95% of both Group VIB (eg, Mo) and Group VB (eg, V) metals from the spent supported catalyst into the solution phase is possible. The filtrate typically contains greater than about 70 wt%, or 75 wt%, or 80 wt%, or 85 wt%, or 90 wt%, or 95 wt% of the Group VIB metal, or greater than about 50 wt%, or 60 wt%, or 70 wt%, or 80 wt%, or 90 wt%, or 95 wt% of the Group VB metal originally present on the deoiled spent supported catalyst, or greater than about 70 wt%, or 75 wt%, or 80 wt%, or 85 wt%, or 90 wt%, or 95 wt% of the Group VIB metal and greater than about 50 wt%, or 60 wt%, or 70 wt%, or 80 wt%, or 90 wt%, or 95 wt% of the Group VB metal originally present on the deoiled spent supported catalyst. Optionally, the filtrate contains potassium molybdate, potassium vanadate, or a mixture thereof.

[0059] The residue from the water leaching / extraction step typically contains Group VB and / or Group VIB and / or Group VIIIB metal oxide solids. The residue is then separated from the filtrate and dried under suitable conditions.

[0060] An illustrative water recovery method or process according to an embodiment of the invention is shown generally in Figure 2. Filtrate (43) from one or more sources, such as a spent supported catalyst filtrate stream (43) comprising an aqueous mixture of Group VIB and Group VB metal compounds from the process shown in Figure 1, is mixed (50) with ammonium salt (52) under metathesis reaction conditions to convert the metal compounds to ammonium Group VB and Group VB metal compounds. The metathesis reaction mixture is then subjected (60) to crystallization conditions effective to crystallize the ammonium Group VB metal compound. The crystallized ammonium Group VB metal compound is then sent (63) for separation (70) and the ammonium Group VB and Group VIB metal compound filtrate (73) is recovered. Optionally, a saturated ammonium Group VB metal compound wash (72) at a predetermined wash temperature may be used to filter and wash the ammonium Group VB metal compound crystals. The ammonium Group VB metal compound is then subjected to heating (80) under conditions effective to release ammonia (74) to remove the ammonia and separately recover the Group VB metal compound (83) and ammonia (84). The ammonium Group VIB metal compound filtrate (73) obtained from the separation step (70) is then sent to mix (90) with an inorganic acid (92) under conditions effective to form a mixture of a Group VIB metal oxide compound precipitate and an ammonium salt of the inorganic acid. The precipitate and salt mixture is then conveyed (93) to separate (100) and recover the Group VIB metal oxide compound precipitate (101). Optionally, an ammonium Group VIB metal oxide compound wash solution (102) at a wash temperature may be used to filter and wash the Group VIB metal oxide compound precipitate (101). The filtrate (103) from separation (100) typically contains <1,000 mg / L Mo, <100 mg / L V, and up to 40,000 mg / L K. +ions, and up to 15,000 mg / L NH4 + The resulting solution contains iron ions and can be subjected to iron precipitation for removal of residual metals (Mo and V). The transition metal depleted solution can then be subjected to cooling crystallization under vacuum to optionally produce a blend of agricultural grade KNO3 and NH4NO3 fertilizer chemicals.

[0061] Mixing (50, FIG. 2) of the filtrate (43, FIG. 2) with an ammonium salt (52, FIG. 2) is typically carried out under conditions effective to convert the Group VIB and VB metal compounds to ammonium Group VB and VB metal compounds. Seed crystals such as ammonium metavanadate (AMV) may be used, typically at a concentration of about 2000-8000 ppm, about 4000-6000 ppm, or about 5000 ppm. When the AMV seeds are introduced, the pH range is typically less than about 8. One skilled in the art can readily determine a suitable method for carrying out the metathesis reaction, but one useful procedure is to first lower the pH to about 9 using an inorganic acid, e.g., nitric acid, followed by introduction of ammonium nitrate at a pH of less than about 8, preferably less than 8, or in the range of 7.5-8.5, or 7.5-8, or 8, and introduction of the AMV seeds. The metathesis reaction conditions can also include a temperature of less than about 80° C., or less than about 70° C., or in the range of about 50-70° C., or 55-65° C., or about 60° C., and / or a reaction time in the range of about 0.25-2 hours, or about 0.25-1.5 hours, or about 0.5-1.5 hours, or about 1-2 hours. In some cases, the metathesis reaction conditions include the sequential steps of adjusting the pH of the aqueous mixture to a range of about 8 to about 9, adding an ammonium salt to the aqueous mixture, and adding an ammonium Group VB metal compound species to the aqueous mixture at a pH in the range of about 7.5-8.5, preferably about 8.

[0062] During mixing of the filtrate and the metathesis reaction taking place, for example, when the filtrate is removed from a spent supported catalyst containing, for example, Mo, Ni, V, Fe, C, and S, the following representative reactions are believed to form soluble (Mo) metal compounds and (V) metal compounds at high temperatures: NH4NO3+KVO3→NH4VO 3↓ +KNO3(18) 2NH4NO3+K2MoO4→(NH4)2MoO4+2KNO3(19)

[0063] The crystallization conditions may typically be reduced temperature and pressure, for example, at about 10° C. under a vacuum of about 21 in. Hg, to produce ammonium metavanadate (AMV) crystals. It will be apparent to one skilled in the art that since AMV solubility decreases with temperature, various temperature and pressure (vacuum) conditions, as well as crystallization times, may be used. Generally, temperatures ranging from above 0° C. to about 15° C. or above 0° C. to about 10° C., vacuum conditions, and crystallization times of about 1 hour to about 6 hours, about 1 hour to about 4 hours, or about 1 hour to about 3 hours are useful. Filtration of the crystals and washing with a wash at low temperature, for example, an AMV wash of about 5000 ppm V at about 10° C., may be used. Multiple washes, about 2 to 5 times or about 3 times, may also be used, with the wash being recycled to the crystallization step. Typically, wash temperatures in the range of from above 0° C. to about 15° C. or from above 0° C. to about 10° C., or a wash solution temperature of about 10° C. have been found to be suitable, preferably the crystallized ammonium group VB metal compound and the wash solution comprise ammonium metavanadate, and optionally the wash solution is recycled for crystallization of the ammonium group VB metal compound. Generally, the ammonium group VB metal compound wash solution comprises the same ammonium group VB metal compound as the crystallized ammonium group VB metal compound, or the ammonium group VB metal compound of the wash solution is the same ammonium group VB metal compound as the crystallized ammonium group VB metal compound.

[0064] The ammonium group VB metal compound may then be treated according to standard procedures, for example, by heating at a temperature in the range of about 200-450°C, about 300-450°C, about 350-425°C, or about 375-425°C for a time sufficient to release ammonia in an amount of at least about 90%, about 95%, about 98%, or about 99% of the amount present in the ammonium group VB metal compound. The group VB metal compound is then further treated, for example, by melting in a smelter and discharging the melt onto a flaker wheel to produce group VB metal compound flakes. The total recovery of the group VB metal present in the group VIB metal compound and the deoiled spent supported catalyst containing the group VB metal compound may be greater than about 85 wt%, greater than about 90 wt%, or greater than about 95 wt%.

[0065] The acidification conditions for contacting the ammonium Group VIB metal compound filtrate with an inorganic acid include introducing the inorganic acid at a temperature in the range of about 50-80° C., about 50-70° C., or about 55-70° C. to adjust the pH to about 1-3, about 1-2, or about 1, and a reaction time of about 1 hour to about 4 hours, or about 1 hour to about 3 hours, or about 1 hour to about 2 hours, and preferably the inorganic acid includes or is nitric acid or sulfuric acid.

[0066] During the acidification reaction, when the filtrate is removed from a spent supported catalyst containing, for example, Mo, Ni, V, Fe, C and S, the following representative reactions are believed to form insoluble (Mo) metal compounds: (NH4)2MoO4+2HNO3+H2O→MoO3·2H2O ↓ +2NH4NO3(20)

[0067] After the acidification reaction, the slurry may be cooled and the liquid and solids may be separated using filtration. The conditions for washing the Group VIB metal oxide compound precipitate may be by reslurrying the filter cake with an ammonium Group VIB metal compound wash at a wash temperature ranging from above 0°C to about 15°C or above 0°C to about 10°C, or at a wash temperature of about 10°C, at a pH of about 1.0, for 15 minutes at 25 wt% solids. Typically, when the spent catalyst contains Mo as the Group VIB metal, the wash solution contains ammonium heptamolybdate (AHM) with a molybdenum depletion at pH 1.0. After refiltering the slurry, the cake may be reslurried two more times with fresh ammonium heptamolybdate solution at pH 1.0 to reduce the K content in the MoO3 cake to less than 0.5 wt%. As with all wash steps, the wash solution may be optionally recycled, for example for washing Group VIB metal oxide compounds. Generally, the ammonium Group VIB metal oxide compound cleaning solution contains the same ammonium Group VIB metal oxide compound as the deposited ammonium Group VIB metal oxide compound, or the ammonium Group VIB metal oxide compound of the cleaning solution is the same ammonium Group VIB metal oxide compound as the deposited ammonium Group VIB metal compound.

[0068] The total recovery of the Group VIB metals present in the Group VIB metal compounds and the deoiled spent supported catalyst containing the Group VB metal compounds may be greater than about 85 wt%, greater than about 90 wt%, or greater than about 95 wt%. Figure 3 shows a process schematic combining the roasting and extraction methods shown in Figure 1 with the water recovery method shown in Figure 2. The above discussion of each of Figures 1 and 2 is directly applicable to the combined process schematic shown in Figure 3. EXAMPLES

[0069] The following examples provide results of metal recovery from spent supported catalyst in accordance with the claimed invention. Results of metal recovery using potassium carbonate (potash) in accordance with an embodiment of the invention are provided along with comparative results in which potassium carbonate is not used. Comparative results using sodium carbonate instead of potassium carbonate are also provided.

[0070] Reference (Comparative) Examples 1A and 1B provide the results of roasting the intact spent supported catalyst followed by potassium hydroxide (caustic potassium) leaching of the calcine. Both fixed bed (FB) and ebullated bed (EB) supported catalysts were used in Examples 1A and 1B.

[0071] Examples 2A and 2B provide the results of roasting the spent supported EB and FB catalyst-sodium carbonate mixtures and hot water leaching of the supported catalyst-sodium carbonate calcine.

[0072] Comparative Examples 3A and 3B provide the results of roasting spent supported EB and FB catalyst-potassium carbonate mixtures and hot water leaching of the supported catalyst-potassium carbonate calcine.

[0073] Example 4A provides crystallization results for each of the leach filtrates from Examples 1B, 2B, and 3B after completion of the metathesis reaction for recovery of the Group VB metal. Example 4B provides results for Group VIB metal recovery after metal precipitation and from each of the remaining Group VB metal-barren liquors of Example 4A.

[0074] A rotary quartz tube furnace was used for controlled batch oxidation of deoiled spent catalyst samples. The tube furnace included three zones, furnace zones 1, 2, and 3 located along the length of the tube furnace, with zone 1 being the feed inlet in the left third of the furnace, zone 2 being the middle third of the center section of the furnace, and zone 3 being the outlet in the right third of the tube furnace. Figures 4-9 provide the heat and off-gas results as described in the Examples. In each of Figures 4-9, the number and letter identifiers are written as follows: [Table 0]

[0075] Example 1A - Roasting of Spent Supported Ebullated Bed (EB) and Fixed Bed (FB) Catalysts (as is): Controlled batch oxidation of up to 260 g of deoiled spent catalyst was performed in a 4 in. diameter x 14 in. working length rotary quartz tube furnace at temperatures between 600°C and 625°C under an air flow of 3 to 5 standard liters per minute with residence times up to 8 hours simulating multiple hearth furnace conditions to obtain calcines containing less than 0.1 wt.% C and less than 0.7 wt.% S, respectively.

[0076] The operation started with a rapid temperature increase to 500°C under argon gas flow to remove residual hydrocarbons in the spent catalyst, followed by a slow temperature increase to an operating bed temperature of 620°C under 3 slpm air flow, and the CO2 and SO2 were removed. x The stepwise temperature control, which consisted of an extended hold time under 4 slpm air with emission measurements followed by slow cooling under excess air flow during reaction quenching, was essential to avoid significant heat release which would result in loss of Mo and sintering of the solid.

[0077] Weight losses of up to 26% (Tables 7 and 8) were observed in the EB and FB spent catalyst calcines, corresponding to almost complete removal of S (<0.7 wt%) and C (<0.06 wt%) along with conversion of metal sulfides to metal oxides.

[0078] Tables 1 and 2 provide metal analysis test results for the roaster spent supported catalyst feedstock and the calcine produced. [Table 1] [Table 2]

[0079] Reactions (1.1) to (1.6) below represent the appropriate combustion reactions. The Gibbs free energies at 600°C suggest an oxidation order of V>Mo>Fe>Ni, and the free energies of CO2 and SO2 at 600°C suggest that C will burn at a faster rate than S. V2S3+11 / 2O2=V2O5+3SO2 ↑ (1.1) ΔG 873°K =-1,585kJ / g.mol MoS2+7 / 2O2=MoO3+2SO2 ↑ (1.2) ΔG 873°K =-879kJ / g.mol 2FeS+7 / 2O2=Fe2O3+2SO2 ↑ (1.3) ΔG 873°K =-484kJ / g.mol NiS+3 / 2O2=NiO+SO2 ↑ (1.4) ΔG 873°K =-375kJ / g.mol C+O2=CO2 ↑ (1.5) ΔG 873°K =-396kJ / g.mol S+O2=SO2 ↑ (1.6) ΔG 873°K =-298kJ / g.mol

[0080] Unexpected phases detected by XRD included nickel vanadate in the as-calcined roaster feed (alkali-free). The absence of NiO and the presence of Ni3(VO4)2 occurred at 500°C-600°C in the presence of air where the following reactions (1.7)-(1.9) were preferred over reaction (1.4) and the reaction spontaneity decreased slightly with increasing temperature. V2S3+3NiS+10O2=Ni3(VO4)2+6SO2 ↑ (1.7) ΔG 773°K =-2,823 kJ / g.mol V2S3+3NiS+10O2=Ni3(VO4)2+6SO2 ↑ (1.8) ΔG 823°K =-2,789 kJ / g.mol V2S3+3NiS+10O2=Ni3(VO4)2+6SO2 ↑ (1.9) ΔG 873°K =-2,754 kJ / g.mol MoO3+NiO=NiMoO4(1.10) ΔG 873°K=-20kJ / g.mol

[0081] These S oxidation reactions were supported by observing the gas composition of the thermal oxidation products of the roasting sequence of EB and FB spent supported catalysts (Figures 4 and 5). Maximum SO2 emissions of up to 3 vol% for EB spent catalyst and up to 6 vol% for RDS spent catalyst were observed within the first hour of oxidation, indicating a combination of reaction progression (1.7)-(1.9). Up to 26% of the total S in the feedstock present in the Ni-V stoichiometry in the described reaction could have been combusted to SO2 within the first hour after hydrocarbon removal under argon at 500 °C. Small amounts of the spinel phase NiMoO4 were also detected in the diffraction patterns (reaction 1.10), and this phase is non-leachable in caustic potassium or caustic soda. The dissolution of nickel vanadate is also minimal in both alkaline and acidic conditions.

[0082] Example 1B - Leaching of calcine with caustic potash (KOH): The calcine was used as is and ground to 100% passing through a 65-Tyler for leaching. Caustic potassium (KOH, 15 wt% solution) leaching of both EB and FB calcine of Example 1A at the two particle size distributions was carried out at 75°C, 15 wt% solids, pH 10.0-11.0, and a 3 hour hold time. The leach residue was vacuum filtered, washed, dried, and submitted for analysis. The leach solution was allowed to settle until close to hydrometallurgical separation of V from Mo (see Examples 4A and 4B).

[0083] Table 3 shows that for EB calcine, the crushed 65-Tyler mesh material mostly resulted in more metal (Mo and V) extraction, whereas FB calcine did not appear to result in more metal extraction with the finer sized 65-Tyler mesh fraction.

[0084] Both the kinetic extraction in Table 3 and the leaching residue in Table 4 show that EB calcination is up to 5% better for Mo and 15% better for V extraction compared to FB calcination. The improved metal extraction from EB calcination spent catalyst can be attributed to the fluidization of the catalyst in the reactor bed due to the superior mixing properties and the frequent replacement of EB spent catalyst in the refinery setup. In comparison, in a static bed configuration such as in a FB reactor, metal extraction may be more difficult due to the accumulation of auxiliary contaminants, longer replacement periods of the FB catalyst, and simultaneous trapping of metals within the spent catalyst material.

[0085] In Table 2, the molar ratios of calcine V / Ni indicate that for both EB and FB calcine, up to 82% leaching would be possible if all Ni present tied up V as non-leachable nickel vanadate and no other encapsulation mechanisms were present. The data in Table 3 suggest that this observation appears to be true for intact EB calcine, whereas V lock-up is considerably higher for intact FB calcine. Table 4 also supports the understanding that, based on the molar V / Ni ratios, the extent of V lock-up as refractory non-leachable nickel vanadate is significantly greater for FB-type spent catalyst compared to EB-type spent catalyst.

[0086] XRD scan of the leach residue confirmed that the insoluble matter was mainly Al2O3, Ni3(VO4)2, Fe2O3 and a small amount of α-NiMoO4. The residual C and S in the leach residue, determined by Leco analysis, was on average less than 0.1 wt%.

[0087] Dissolution of up to 5% of the total Al present in the calcine was observed at alkaline potassium pH. Dissolution of up to 30% of Fe was also detected under alkaline conditions. The polynuclear complex Fe(OH)4, present at pH > 10.0 -1 is thought to have contributed to the solubility of iron.

[0088] For both EB and FB ground calcine, it was observed that the EB and FB calcine mass was dissolved in KOH by up to 50% (Tables 5 and 6), with the remaining mass constituting the insoluble matter in the washed leach residue.

[0089] Tables 7 and 8 provide theoretical metal mass balances per unit operation for both EB and FB prototypes. As noted, EB Mo and V extractions were up to 5% and 15%, respectively, higher than the corresponding FB Mo and V extractions. Both Mo and V extractions were approximately 5% higher for the ground EB calcine versus the neat calcine, with only minor differences observed for the ground versus the neat FB calcine. All subsequent testing used calcine ground to 100% passing a 65 Tyler mesh. [Table 3] [Table 4] [Table 5] [Table 6] [Table 7] [Table 8]

[0090] From the foregoing, note the following about EB and FB metal extraction: For alkali-free ebullated bed (EB) spent catalyst, pyrometallurgical extraction rates of Mo and V of up to 85% and up to 81%, respectively, and hydrometallurgical recoveries of Mo and V of up to 99% and up to 95%; total metal recoveries are predicted to be 84% Mo and 77% V. For alkali-free fixed-bed (FB) spent catalyst, pyrometallurgical extraction rates of up to 80% Mo and up to 65% V, and hydrometallurgical recovery rates of up to 99% Mo and up to 95% V; total metal recoveries are predicted to be 79% Mo and 62% V.

[0091] Example 2A - Roasting of Spent EB and FB Supported Catalysts with Potassium Carbonate and Hot Water Leaching of KCO / Spent Supported Catalyst Calcines: The relatively low Mo and V extractables obtained from KOH leaching of intact alkali-free roasted spent catalyst were of concern from the perspective of commercial metal recovery and project economics. Potassium carbonate was stoichiometrically blended with the roaster feedstock to define the metal liberation mechanism. The objective was to increase the metal extraction yield by removing non-leachable species such as nickel vanadate and nickel molybdate.

[0092] The following reactions (2.1)-(2.9) represent the oxidation reactions of metals with K2CO3; the Gibbs free energies at 600°C suggest the order of favorable oxidation is V>Mo>Fe>Ni>C>S. The free energies at 600°C with CO2 and SO2 suggest that C burns at a faster rate than S. MoS2+3K2CO3+9 / 2O2=K2MoO4+2K2SO4+3CO2 ↑ (2.1)ΔG 873°K =-1,571 kJ / g.mol V2S3+4K2CO3+7O2=2KVO3+3K2SO4+4CO2 ↑ (2.2)ΔG 873°K =-2,600kJ / g.mol NiS+K2CO3+2O2=NiO+K2SO4+CO2 ↑ (2.3)ΔG 873°K =-655kJ / g.mol 2FeS+2K2CO3+9 / 2O2=Fe2O3+2K2SO4+2CO2 ↑ (2.4)ΔG 873°K =-764kJ / g.mol C+O2=CO2 ↑(2.5)ΔG 873°K =-396kJ / g.mol S+O2=SO2 ↑ (2.6)ΔG 873°K =-298kJ / g.mol K2CO3+SO2+1 / 2O2=K2SO4+CO2 ↑ (2.7)ΔG 873°K =-280kJ / g.mol NiMoO4+K2CO3=K2MoO4+NiO+CO2 ↑ (2.8)ΔG 873°K =-111kJ / g.mol Ni3(VO4)2+K2CO3=2KVO3+3NiO+CO2 ↑ (2.9)ΔG 873°K =-130kJ / g.mol

[0093] When the refractory and non-leachable NiMoO4 is formed via the interaction of the intermediate MoO3 with NiO species (reaction 1.10), the negative free energy indicates that the compound is converted to mostly water-soluble molybdate in the presence of alkali (reaction 2.8).

[0094] Without being bound by theoretical considerations, the negative Gibbs free energy of reaction (2.9) can suggest a degree of spontaneity of the forward reaction in the presence of potassium. Based on reactions (1.7)-(1.9), the formation of non-leachable nickel vanadate also seems certain at 500°C-600°C+, considering the higher free energy compared to reaction (2.2). This is further supported by the observation of 2 hours of SO2 evolution following the first hour of hydrocarbon removal under argon (Figures 6 and 7). Furthermore, the presence of SO2 under alkaline roasting implies that reactions (1.7)-(1.9) were predominant. Reaction (2.9) indicates that all or most of the refractory nickel vanadate would be converted to water-leachable potassium vanadate moieties.

[0095] Controlled batch oxidation was carried out on up to 260 g of deoiled spent catalyst blended with various stoichiometric additions of K2CO3 (Rocky Mountain Reagents, 28% passing 300 μm) in a 4 in. diameter x 14 in. working length rotary quartz tube furnace under air flow of 3-5 standard liters per minute. Temperatures were maintained between 600°C and 625°C with residence times up to 8 hours to simulate multiple hearth furnace conditions, producing calcine with minimal petroleum coke and S combustion to soluble sulfates.

[0096] Both deoiled spent catalysts (EB and FB) were thoroughly blended with anhydrous K2CO3 at various stoichiometric Mo and V content levels in the spent materials, i.e., a) stoichiometric, b) 15% above stoichiometric, and c) 25% above stoichiometric.

[0097] The operation started with a rapid temperature increase to 500°C under argon gas flow to remove residual hydrocarbons in the spent catalyst, followed by a slow temperature increase to an operating bed temperature of 620°C under 3 slpm air flow, and the CO2 and SO2 were removed. x The stepwise temperature control, with extended hold times up to 4 slpm air with emission measurements, followed by slow cooling under excess air flow during reaction quenching, was essential to avoid significant heat release that would result in loss of Mo and sintering of the solid.

[0098] Weight losses of up to 78% (Tables 12 and 13) were observed in the EB and FB spent catalyst calcines, which correspond to the conversion of metal sulfides to metal oxides along with the conversion of alkali and coke to CO2. Gas evolution of sulfur dioxide was evident over a 2-hour period after air entry was started, indicating that reactions (1.7)-(1.9) occurred or the conversion of vanadium and nickel sulfides to refractory nickel vanadate in the presence of air. Subsequent conversion of nickel vanadate to water-leachable alkali vanadate is indicated per reaction (2.9).

[0099] These S oxidation and stoichiometric alkali fixation reactions after hydrocarbon removal under argon are supported by observing the gas composition of the thermal oxidation products during the roasting sequence of EB and FB spent catalysts (Figs. 6 and 7).

[0100] Table 9 presents calcine assays with varying potassium content for both types of spent catalysts, with the high C and S levels mainly due to the formation of residues K2CO3 and K2SO4 from reactions (1)-(4) and (7)-(9). These residues reduce the Mo, Ni, and V contents by mass in the calcine, but the molar ratio of V / Ni is close to the similar ratios in Table 2. [Table 9]

[0101] Example 2B - Hot Water Leaching of Potassium Carbonate Calcine: K2CO3 calcine was leached in hot water at 75°C (pH 10.5-11.0) at 15 wt% solids for 3.0 h without changing the pH of the sample. The leach residue was vacuum filtered, washed, dried, and submitted for analysis. The leach solution was allowed to settle until close to hydrometallurgical separation of V from Mo, as shown in Examples 4A and 4B.

[0102] It was observed that in both ground calcine K2CO3 roasted products, EB and FB calcine lumps were dissolved in hot water up to 78% (Tables 12 and 13), while the remaining lumps constituted the insoluble matter in the washed leach residue.

[0103] EB calcine From the hot water leaching of potassium calcine, extraction of Mo and V up to 91% each (Table 10) was achieved, with a weight loss of up to 77% evident (Table 12). Metal analysis test results of the leach residue are given in Table 11, with Al and Ni identified as constituting 40 wt% and 6.5 wt% of the unreacted solid phase, respectively.

[0104] XRD scans of the leach residues identified the insolubles as mainly Al2O3, NiO, some Ni3(VO4)2, Fe2O3 and small amounts of α-NiMoO4. The C and S residues in the leach residues, as determined by Leco analysis, averaged less than 1 wt%.

[0105] The data also reveal that, within experimental and assay discrepancies, a stoichiometric amount of potassium is sufficient for optimal metal extraction.

[0106] FB Calsign From the hot water leaching of potassium calcine, extraction of Mo and V up to 91% and 88%, respectively (Table 10), was achieved with a weight loss of up to 78% (Table 13). Metal analysis test results of the leach residue are given in Table 11, with Al and Ni identified as constituting 40 wt% and 7.4 wt%, respectively, of the unreacted solid phase.

[0107] XRD scans of the leach residues identified the insolubles as mainly Al2O3, NiO, some Ni3(VO4)2, Fe2O3 and small amounts of α-NiMoO4. The C and S residues in the leach residues, as determined by Leco analysis, averaged less than 1 wt%.

[0108] Table 11 also supports what was stated earlier that the degree of V lock-up as refractory unreacted nickel vanadate is greater in the FB type K2CO3 roasted spent catalyst compared to the EB type K2CO3 roasted spent catalyst, based on the molar V / Ni ratio.

[0109] The data also show that, within experimental and assay discrepancies, a stoichiometric amount of potassium is sufficient for optimal metal extraction. [Table 10] [Table 11] [Table 12] [Table 13]

[0110] From the foregoing, note the following about EB and FB metal extracts: For ebullated bed (EB) spent catalyst, pyrometallurgical extraction rates of Mo and V of up to 90% and hydrometallurgical recovery rates of Mo and V of up to 99% and 95% respectively; total metal recoveries are predicted to be 89% Mo and 87% V. For fixed bed (FB) spent catalyst, pyrometallurgical extraction rates of Mo up to 91% and V up to 87%, and hydrometallurgical recovery rates of Mo up to 99% and V up to 95%; total metal recoveries are predicted to be 90% Mo and 83% V.

[0111] Comparative Example 3A - Roasting of EB and FB Spent Supported Catalysts with Sodium Carbonate and Hot Water Leaching of NaCO / Spent Supported Catalyst Calcines: Soda ash or Na2CO3 may be used commercially during roasting of spent catalysts where metal recovery is supported by hydrometallurgical circuits to produce various Mo and V chemicals for catalyst and chemical manufacturing. Based on favorable thermodynamic assessment of reaction spontaneity, sodium carbonate was stoichiometrically blended with the roaster feedstock to define the metal liberation mechanism. The objective was to evaluate metal extraction along with K2CO3 roasting.

[0112] The following reactions (3.1)-(3.9) represent the oxidation reactions of metals with Na2CO3. The Gibbs free energies at 600°C suggest the order of favorable oxidation is V>Mo>Fe>Ni>C>S. The free energies at 600°C with CO2 and SO2 suggest that C burns at a faster rate than S. MoS2+3Na2CO3+9 / 2O2=Na2MoO4+2Na2SO4+3CO2 ↑ (3.1)ΔG 873°K =-1,504kJ / g.mol V2S3+4Na2CO3+7O2=2NaVO3+3Na2SO4+4CO2 ↑ (3.2)ΔG 873°K =-2,506 kJ / g.mol NiS+Na2CO3+2O2=NiO+Na2SO4+CO2 ↑ (3.3)ΔG 873°K =-630kJ / g.mol 2FeS+2Na2CO3+9 / 2O2=Fe2O3+2Na2SO4+2CO2 ↑ (3.4)ΔG 873°K =-739kJ / g.mol C+O2=CO2 ↑ (3.5)ΔG 873°K =-396kJ / g.mol S+O2=SO2 ↑ (3.6)ΔG 873°K =-298kJ / g.mol Na2CO3+SO2+1 / 2O2=Na2SO4+CO2 ↑ (3.7) ΔG 873°K =-255kJ / g.mol NiMoO4+Na2CO3=Na2MoO4+NiO+CO2 ↑ (3.8)ΔG 873°K =-96kJ / g.mol Ni3(VO4)2+Na2CO3=2NaVO3+3NiO+CO2 ↑ (3.9)ΔG 873°K =-113kJ / g.mol

[0113] When the refractory and non-leachable NiMoO4 is formed via the interaction of the intermediate MoO3 with NiO species (reaction 1.10), the negative free energy in reaction (3.8) indicates that the compound is converted to mostly water-soluble molybdate in the presence of alkali.

[0114] The negative Gibbs free energy of reaction (3.9) implies the degree of spontaneity of the forward reaction in the presence of soda ash, and based on reactions (1.7)-(1.9), the formation of non-leachable nickel vanadate seems certain at 500°C-600°C+, considering the higher free energy compared to reaction (3.2). However, this was not revealed by SO2 evolution during observation of the thermal oxidation product gas composition following the first hour of hydrocarbon removal under argon (Figures 8 and 9). The negligible presence of SO2 under soda ash roasting could indicate that soda ash is more effective than potassium in capturing and neutralizing SO2 during gas transport in the solid phase of each reaction (3.7). Reaction (3.9) suggests that all or most of the refractory nickel vanadate would be converted to water-leachable potassium vanadate moieties.

[0115] Controlled batch oxidation was carried out on up to 260 g of deoiled spent catalyst blended with various stoichiometric additions of anhydrous Na2CO3 (Rocky Mountain Reagents, 28% passing 300 μm) in a 4 in. diameter x 14 in. working length rotary quartz tube furnace under air flow of 3-5 standard liters per minute. Temperatures were maintained between 600°C and 625°C with residence times up to 8 hours to simulate multiple hearth furnace conditions, producing calcine with minimal petroleum coke and S combustion to soluble sulfates.

[0116] Both deoiled spent supported catalysts (EB and FB) were thoroughly blended with anhydrous Na2CO3 at various stoichiometric Mo and V content levels in the spent materials, i.e., a) stoichiometric, and b) 25% above stoichiometric.

[0117] The operation started with a rapid temperature increase to 500°C under argon gas flow to remove residual hydrocarbons in the spent catalyst, followed by a slow temperature increase to an operating bed temperature of 620°C under 3 slpm air flow, and the CO2 and SO2 were removed. xAn extended hold time of up to 4 slpm air with emission measurements followed by slow cooling under excess air flow during reaction quenching was used in order to avoid solid fusibility, sintering and significant heat release which would result in loss of Mo.

[0118] A maximum weight loss of 78% (Tables 17 and 18) was observed in the EB and FB spent catalyst calcine, which corresponds to the conversion of metal sulfides to metal oxides along with the conversion of alkali and coke to CO2. After the air flow was started, minimal SO x Evolution was evident throughout the run, indicating that evolved SO2 in the presence of soda ash was directly converted to sulfate.

[0119] These S oxidation and stoichiometric alkali fixation reactions after hydrocarbon removal under argon were supported during the observation of the gas composition of the thermal oxidation products for the roasting sequences of EB and FB spent catalysts (Figs. 8 and 9).

[0120] Table 14 presents calcine assays with varying potassium content for both types of spent supported catalysts. The high C and S levels are mainly due to the formation of residual K2CO3 and K2SO4 in reactions (3.1)-(3.4) and (3.7), respectively. These residues reduce the Mo, Ni, and V content by mass in the calcine, but the molar ratio of V / Ni is close to the similar ratio in Table 2. [Table 14]

[0121] Example 3B - Hot Water Leaching of Sodium Carbonate Calcine: Na2CO3 calcine was leached in hot water at 75°C (pH 10.5-11.0) at 15 wt% solids for 3.0 h without changing the pH of the sample. The leach residue was vacuum filtered, washed, dried, and submitted for analysis. The leach solution was allowed to settle until close to hydrometallurgical separation of V from Mo, as exemplified in Examples 4A and 4B.

[0122] It was observed that in both ground calcine Na2CO3 roasted products, EB and FB calcine lumps were dissolved in hot water up to 78% (Tables 17 and 18), while the remaining lumps constituted the insoluble matter in the washed leach residue.

[0123] EB calcine From the hot water leaching of soda ash calcine, extraction of Mo and V, respectively, up to 88% (Table 15) was achieved, with weight losses of up to 78% evident (Table 17). Metal analysis test results of the leach residue are given in Table 16, with Al and Ni identified as constituting 38 wt% and 7 wt%, respectively, of the unreacted solid phase.

[0124] XRD scan of the leach residue confirmed that the insoluble matter was mainly Al2O3, Ni3(VO4)2, Fe2O3 and a small amount of α-NiMoO4. No NiO residue, the product of reactions (3.3) and (3.9), was detected. The C and S remaining in the leach residue, determined by Leco analysis, was on average less than 0.2 wt%.

[0125] The data also reveal that, within experimental and assay discrepancies, stoichiometric amounts of soda ash are sufficient for optimal metal extraction.

[0126] FB Calsign From the hot water leaching of soda ash calcine, extraction of Mo and V up to 90% and 78%, respectively (Table 15), was achieved with a weight loss of up to 75% (Table 18). Metal analysis test results of the leach residue are shown in Table 16, with Al and Ni identified as constituting 35 wt% and 7.2 wt%, respectively, of the unreacted solid phase.

[0127] XRD scans of the leach residues confirmed that the insoluble matter was mainly Al2O3, Ni3(VO4)2, Fe2O3 and small amounts of α-NiMoO4. NiO residues, the product of reactions (3.3) and (3.9), were again not detected. C and S remaining in the leach residues, as determined by Leco analysis, averaged less than 0.2 wt%.

[0128] Extraction of V in hot water after soda ash roasting of FB spent catalyst shows significantly lower yields compared to its EB spent catalyst. Comparing the V / Ni molar ratios in Table 11 and Table 16 (hot water leaching residue), there is a significant lock-up of V in Ni, suggesting that soda ash roasting was unable to liberate the metal from the refractory nickel vanadate. This is especially true for FB calcine.

[0129] The data also show that, within experimental and assay discrepancies, stoichiometric amounts of soda ash are sufficient for optimal metal extraction. [Table 15] [Table 16] [Table 17] [Table 18]

[0130] The metal extraction and total metal recovery results described in the examples (FIGS. 4-9) are summarized in Table 19. [Table 19]

[0131] From the foregoing, note the following about EB and FB metal extraction: For ebullated bed (EB) spent catalyst, pyrometallurgical extraction rates of up to 89% Mo and up to 88% V, and hydrometallurgical recovery rates of up to 99% Mo and up to 95% V. Total metal recoveries are predicted to be 88% Mo and 84% V.

[0132] For fixed bed (FB) spent catalyst, pyrometallurgical extraction rates of up to 90% Mo and up to 78% V, and hydrometallurgical recovery rates of up to 99% Mo and up to 95% V. Total metal recoveries are predicted to be 89% Mo and 74% V.

[0133] General conclusions for the torrefaction scenario: The total metal recoveries from two types of spent supported catalysts under various roasting scenarios are shown in Table 20.

[0134] Regarding Examples 1, 2 and 3, the following should be noted. a) The recovery rates of Mo and V in EB roasted with both alkalis are increased by up to 5% and 10%, respectively, compared to those without. b) The recovery rates of Mo and V in FB roasted with both alkalis are increased by up to 10% and up to 20%, respectively, compared to without. c) The Mo recovery rates from both EB and FB types using either alkali were comparable. d) The V recovery rate from EB spent catalyst using potassium roasting was up to 3% better than that of soda ash roasting; and e) The V recovery difference from FB spent catalyst used with potash roasting was up to 9% more than soda ash roasting. [Table 20]

[0135] Example 4A - Crystallization of Ammonium Metavanadate from Leach Solution: A stirred solution of the leach filtrate (pH 10.0 or greater) of Examples 1B, 2B, and 3B was heated to 60° C. Sufficient 70% concentrated HNO3 acid was added to lower the pH to about 8.8.

[0136] 100 gpL of NH4NO3 crystals was added and the pH was adjusted to approximately 7.5 with HNO3 or NH4OH.

[0137] If the vanadium concentration of the solution was less than 10 gpL, 10 gpL of AMV seed / spike was added in powder form to the hot, stirred solution.

[0138] The metathesis reaction was continued at 60° C. for 1.5 hours while maintaining the pH at 7.0-8.0.

[0139] The double substitution shown in Reactions 4.1 and 4.2 is NH4 + and K. + or Na + This constitutes a metathesis or ion exchange between NH4NO3+KVO3=NH4VO 3↓ +KNO3(4.1) 2NH4NO3+K2MoO4=(NH4)2MoO4+2KNO3(4.2)

[0140] The solution was then transferred to a vacuum cooled crystallizer under gentle rotation at 10° C. and 21 inches Hg for 3 hours while crystallization continued.

[0141] The AMV crystals were vacuum filtered and the filtrate was set aside for Mo precipitation; the crystals were washed with three pore volumes of 4,800 mg / L pure AMV solution cooled to 10° C.; the wash solution may be reused until the residual Mo concentration is increased up to 25,000 ppmw, after which the wash solution may be recycled to the metathesis circuit.

[0142] The yellowish AMV crystals were dried at 60° C. to 70° C. Table 21 shows that continuous cooling crystallization at 10° C. reduces the V content in the barren liquor.

[0143] The estimated AMV purity is up to 97 wt% NH4VO3, with the remainder being NO3 - As species of Mo and K or Na combined with anions.

[0144] The barren liquor or Mo filtrate was transferred to an acid precipitation circuit for Mo recovery. [Table 21]

[0145] Example 4B - Precipitation of AMV Barren Solution with Molybdenum Trioxide: The stirred barren liquor from the V crystallization circuit was heated to 65°C and then 70% concentrated HNO3 acid was carefully added to bring the pH to approximately 1.0. The pH and temperature were maintained for up to 4 hours with good stirring. The precipitation effect is shown in reaction 4.3.

[0146] (NH4)2MoO4+2HNO3=MoO3.H2O ↓ +2NH4NO3(4.3) Table 22 shows that the lower pH and temperature, and higher HNO3 acid dosage resulted in up to 99% Mo recovery in 2 hours.

[0147] Upon cessation of the reaction, the slurry was cooled to near ambient and filtered. <1,000 mg / L Mo, <100 mg / L V, max 40,000 mg / L K + Ion or Na + ions, and up to 15,000 mg / L NH4 + The barren solution containing the ions was transferred to iron precipitation to remove residual metals. The transition metal-depleted solution was then subjected to cooling crystallization under vacuum to produce a mixture of KNO3 and NH4NO3.

[0148] The cake was washed with 2 pore volumes (PV) pH 1 ambient AHM* and the wash filtrate was recycled. The cake solids were then reslurried at ambient temperature in pH 1 AHM at 25 wt% solids for 15 minutes with stirring.

[0149] The slurry was refiltered and the resulting barren liquor was directed to wash recycle. The filter cake was washed with 4 PV pH 1 ambient AHM and the barren liquor was recycled as wash. The solids were dried at 70°C-100°C.

[0150] The estimated MoO3 purity is ~95 wt% MoO3.H2O, ~0.75 wt% total K or Na and V, and the remainder NH4+ Ions and NO3 - Contains ions.

[0151] Using the washing steps in the order described, the MoO3 product was treated with K + or Na + The alkali metal(s) act as poisons during catalyst synthesis, so it is desirable to lower the value. + or Na + Ion levels may be increased by up to 20% and remain unchanged non-removable K + or Na + The ionic moieties replace hydronium ions in the layered MoO3 structure. Note: *AHM at pH 1 was prepared by acidifying 200 gpL of pure ammonium heptamolybdate (AHM) solution with concentrated HNO3 acid to pH 1 for 2.5 hours at 65 °C. After separation of the liquid and solid, the MoO3 solid may be collected as the final product and the filtrate used as a wash for the commercial MoO3 cake. [Table 22]

[0152] Appendix A - Publications List of patent or literature references that may be relevant to this disclosure: 1.Marcantonio, PJ, “Process for Metals Recovery from Spent Catalyst”, US Patent7,485,267, issued Feb.2009; 2.Bhaduri, RS, “Process for recovering base metals from spent hydroprocessing catalyst”, US Patent7,658,895,issued Feb.9,2010; 3.Marcantonio,P.M.,“Process for Metals Recovery from Spent Catalyst”,US Patent7,736,607,Issued June 15,2010; 4.Bhaduri,R.S.,“Process for recovering base metals from spent hydroprocessing catalyst”,US Patent7,837,960,issued Nov.23,2010; 5.Bhaduri,R.S.,“Process for recovering base metals from spent hydroprocessing catalyst”,US Patent7,846,404,issued Dec.7,2010; 6.Bhaduri,R.S.,“Systems and Methods for Producing a Crude Product」,US Patent No8,236,169,issued Aug.7,2012; 7.Bhaduri,R.S.,“Process for Recovering Metals from Coal Liquefaction Residue containing Spent Catalysts”,US Patent No8,628,735, issued Jan.14,2014; 8.Bhaduri,R.S.,“Hydroprocessing Catalysts and Methods for Making Thereof”,US Patent No8,778,828,issued Jul.15,2014; 9.Bhaduri,R.S.,“Recovery of Vanadium from Petroleum Coke slurry containing solubilized Base Metals”,US Patent No8,815,185, issued Aug.26,2014; 10.Bhaduri,R.S.,“Process for Separating and Recovering Metals”,US Patent No8,815,184, issued Aug.26,2014; 11.Bhaduri,R.S.,“Process for Separating and Recovering Metals”,US Patent No.9,809,870, issued Nov.7,2017; 12.Marafi,M.,Stanislaus,and A.,“Options and Processes for Spent Catalyst Handling and Utilization”,Journal of Hazardous Materials,B101(2003),pp.123-132; 13.Llanos,Z. R.,and Deering,W.G.,“Processes for the Recovery of Metals from Spent Hydroprocessing Catalysts”,Third International Symposium on Recycling of Metals and Engineered Materials,Edited by P.B.Queneau and R.D.Peterson,TMMS,1995,pp425; 14.Llanos,Z.R.,Lacave,J.,and Deering,W.G.,“Treatment of Spent Hydroprocessing Catalysts at Gulf Chemical and Metallurgical Corporation”,SME Annual Meeting,Mar.1986,New Orleans,Louisiana; 15.Marafi M.,Rana,M.S.,and Al-Sheeha H.,“The Recovery of Valuable Metals and Recycling of Alumina from a Waste Spent Hydroprocessing Catalyst:Extraction with Na Salts”,2014,WIT Transactions on Ecology and The Environment,Vol 180,pp15-27; 16.B.B.Kar,P.Datta,and V.N.Misra,“Spent catalyst:secondary source for molybdenum recovery”,Hydrometallurgy,72(2004)87-92,26 May 2003; 17.Furimsky E.,and Massoth F.E.,“Deactivation of hydroprocessing catalysts”,Catalysis Today,1999;52,pp.381-495; 18.C.Bartholomew,J.Birtill,and A.Hassan,“Catalyst Deactivation,Regeneration and Disposal”,The Catalyst Group Resources,Spring House,PA; 19.“Qualitative Inorganic Analysis”,Vogel,A.I.,4th Edition,1953; 20. “Analytical Chemistry”, Treadwell, FP, 5th Edition, 1919; 21. Ullmann's Encyclopedia of Industrial Chemistry, Fifth Edition, Editors: Elvers, B., Hawkins, S., Schulz, G., Volume A16; 22. A. Roine, HSC Chemistry® [Software], Outotec, Pori 2018. The software is available at www.outotec.com / HSC. 23. M. Morishita and A. Navrotsky, “Calorimetric Study of Nickel Molybdate: Heat Capacity, Enthalpy and Gibbs Energy of Formation”, J.Am.Ceramic Society,86

[11] ,pp1927-32,2003.

[0153] Further details regarding the invention and the scope of the present disclosure can be found from the appended claims.

[0154] The above description of one or more embodiments of the invention is primarily for purposes of illustration, and it is recognized that variations may be employed and are incorporated within the spirit of the invention. Reference should be made to the following claims in determining the scope of the invention.

[0155] For purposes of United States patent practice, and where permitted, in other patent offices, any patents and publications cited in the above description of the invention are hereby incorporated by reference to the extent that any information contained therein is consistent with and / or supplements the above disclosure.

Claims

1. 1. A method for recovering metals from a deoiled spent supported catalyst, the catalyst comprising a Group VIB metal, a Group VIIIB metal, a Group VB metal, and a catalyst support material, the method comprising: combining a deoiled spent supported catalyst comprising a Group VIIIB metal, a Group VIB metal, and a Group VB metal with potassium carbonate to form a deoiled spent catalyst / potassium carbonate mixture; heating the deoiled spent supported catalyst / potassium carbonate mixture under oxidizing conditions to reduce the levels of sulfur and carbon and form a spent supported catalyst / potassium carbonate calcine comprising water soluble Group VIB metal compounds, water soluble Group VB metal compounds, and water insoluble Group VIIIB metal compounds; combining the spent supported catalyst / potassium carbonate calcine with water under slurry leaching process conditions to form a spent supported catalyst / potassium carbonate calcine slurry and leaching the water soluble Group VIB metal compound and the water soluble Group VB metal compound from the spent supported catalyst / potassium carbonate calcine; separating and removing a filtrate and a solid residue from the spent supported catalyst / potassium carbonate calcine slurry, the filtrate comprising the water-soluble Group VIB metal compound and the water-soluble Group VB metal compound, and the solid residue comprising a water-insoluble Group VIIIB metal compound; recovering said water-soluble Group VIB metal compound and said water-soluble Group VB metal compound from said filtrate of said spent supported catalyst / potassium carbonate calcine slurry.

2. 2. The method of claim 1, wherein the deoiled spent supported catalyst is substantially free of residual hydrocarbons, free of residual hydrocarbons, or contains residual hydrocarbons in an amount less than about 1000 ppm, less than 500 ppm, or less than 100 ppm.

3. 2. The method of claim 1, wherein the deoiled spent supported catalyst comprises residual hydrocarbons, and the process further comprises heating the catalyst via pyrolysis, optionally under non-oxidizing conditions, sufficient to reduce the level of residual hydrocarbons to an amount less than about 1000 ppm, less than 500 ppm, or less than 100 ppm.

4. 2. The method of claim 1, wherein the oxidation conditions include a temperature in the range of about 575°C to 600°C, 600 to 625°C, or 625 to 650°C.

5. 10. The method of claim 1, wherein the catalyst support material comprises alumina, silica, titania or a combination thereof, or a catalyst support material comprising alumina, silica, titania or a combination thereof is used to prepare the catalyst.

6. The method of claim 1 , wherein the spent supported catalyst does not include or is not a slurry catalyst.

7. The method of claim 1 , wherein the spent supported catalyst is an ebullated bed catalyst or a fixed bed catalyst.

8. 2. The method of claim 1, wherein the oxidative heating conditions comprise heating the deoiled spent supported catalyst at a first temperature in the presence of air or a gas mixture containing up to about 20 vol.% oxygen.

9. Sulfur and carbon levels, individually or together, ２ and S.O. ２ 13. The method of claim 1, wherein the concentration of fluorine is reduced to less than a predetermined amount of less than about 1 wt%, less than about 0.8 wt%, less than about 0.5 wt%, less than about 0.2 wt%, or less than about 0.1 wt%, as measured by thermal oxidation product gas analysis.

10. 2. The method of claim 1, wherein the heating of the spent supported deoiled catalyst / potassium carbonate mixture under oxidizing conditions is in the range of about 600°C to 650°C, about 610°C to 650°C, or about 610°C to 630°C, or at a temperature greater than about 600°C, greater than about 610°C, greater than about 620°C, greater than about 630°C, greater than about 640°C, or greater than about 650°C.

11. 11. The method of claim 10, wherein said heating under oxidizing conditions comprises heating for a time period ranging from about 0.5 to 12 hours, 1 to 8 hours, or 4 to 8 hours, or less than about 12 hours, less than 10 hours, or less than 8 hours.

12. 12. The method of claim 10 or claim 11, wherein the heating comprises initially heating under inert gas heating conditions at a temperature of about 600° C. or 550° C., or less than 500° C. or 450° C. for a period of time ranging from about 0.5 to 4 hours, or 1 to 3 hours, or less than about 4 hours, or less than 2 hours.

13. The heating under oxidizing conditions produces a gas, the thermal oxidation product gas having at least about 20 wt. % O. ２ , less than about 0.25 wt.% CO ２ and less than about 0.25 wt. % SO ２ The method according to any one of claims 10 to 12, wherein the method is carried out so as to include:

14. 2. The method of claim 1, wherein the gas flow conditions during heating of the deoiled spent supported catalyst / potassium carbonate mixture under oxidizing conditions include air and are sufficient to remove any thermal oxidation product gases.

15. 2. The method of claim 1, wherein the slurry leaching process conditions comprise a leaching temperature in the range of about 60-90°C, 60-80°C or 70-80°C, or the leaching temperature is greater than about 60°C or greater than 70°C.

16. 10. The method of claim 1, wherein the slurry leaching process conditions comprise a leaching time in the range of about 1 to 5 hours, about 2 to 5 hours, or about 2 to 4 hours.

17. 2. The method of claim 1, wherein the slurry leaching process conditions comprise a leaching pH in the range of about 9.5 to 11, about 10 to 11, or about 10 to 10.

5.

18. 2. The method of claim 1, wherein the slurry leaching of the potassium carbonate calcine is carried out without modifying the pH.

19. 2. The method of claim 1, wherein the filtrate comprises soluble molybdate or vanadate compounds or a mixture thereof.

20. 2. The method of claim 1, wherein the filtrate comprises more than about 70 wt%, or 75 wt%, or 80 wt%, or 85 wt%, or 90 wt% of the Group VIB metal present in the deoiled spent supported catalyst, or more than about 50 wt%, or 60 wt%, or 70 wt%, or 80 wt%, or 90 wt% of the Group VB metal present in the deoiled spent supported catalyst, or both more than about 70 wt%, or 75 wt%, or 80 wt%, or 85 wt%, or 90 wt% of the Group VIB metal present in the deoiled spent supported catalyst and more than about 50 wt%, or 60 wt%, or 70 wt%, or 80 wt%, or 90 wt% of the Group VB metal present in the deoiled spent supported catalyst.

21. The method of claim 1 , wherein the solid residue comprises a compound solid of a Group VB metal and / or a Group VIB metal and / or a Group VIIIB metal.

22. 10. The method of claim 1, wherein the filtrate comprises potassium molybdate, potassium vanadate, or a mixture thereof.

23. 2. The method of claim 1, wherein the extraction rate of the Group VB metal present in the deoiled spent supported catalyst is greater than about 80 wt%, greater than about 85 wt%, or greater than about 90 wt%.

24. 2. The method of claim 1, wherein the extraction of the Group VIB metal present in the deoiled spent supported catalyst is greater than about 80 wt%, greater than about 85 wt%, or greater than about 90 wt%.

25. wherein said recovering said water-soluble Group VIB metal compound and said water-soluble Group VB metal compound from said potassium carbonate calcine slurry filtrate is carried out by separately recovering said Group VIB metal compound and said Group VB metal compound from said filtrate, said process comprising: contacting the filtrate containing the Group VIB metal compound and the Group VB metal compound with an ammonium salt to form a mixture under metathesis reaction conditions effective to convert the metal compounds to an ammonium Group VB metal compound and an ammonium Group VIB metal compound; subjecting said mixture containing said ammonium Group VB metal compound to conditions effective to crystallize said ammonium Group VB metal compound; filtering and washing the crystallized ammonium Group VB metal compound with a saturated ammonium Group VB metal compound washing solution at a predetermined washing temperature, and separately recovering the ammonium Group VB metal compound and the ammonium Group VIB metal compound filtrate; heating the ammonium Group VB metal compound under conditions effective to release ammonia and separately recovering the Group VB metal compound and ammonia; contacting the ammonium Group VIB metal compound filtrate with the inorganic acid under conditions effective to form a Group VIB metal oxide compound precipitate and an ammonium salt of the inorganic acid; and filtering and washing the Group VIB metal oxide compound precipitate with an ammonium Group VIB metal oxide compound wash solution at a predetermined wash temperature to recover the Group VIB metal oxide compound precipitate.

26. 26. The method of claim 25, wherein the Group VB metal comprises vanadium and / or the Group VIB metal comprises molybdenum.

27. 27. The method of claim 25 or 26, wherein the aqueous mixture comprising the Group VIB metal compound and the Group VB metal compound comprises a potassium salt of the Group VIB compound and a potassium salt of the Group VB metal compound.

28. 28. The method of any one of claims 25 to 27, wherein the ammonium salt contacted with the Group VIB metal compound and the filtrate containing the Group VB metal compound comprises ammonium nitrate.

29. 29. The method of any one of claims 25 to 28, wherein the metathesis reaction conditions comprise a pH in the range of less than about 9, less than about 8.5, in the range of about 7-8.5 or about 8, a temperature in the range of less than about 80°C, less than about 70°C, about 50-70°C, 55-65°C or about 60°C, and / or a reaction time in the range of about 0.25-2 hours, about 0.25-1.5 hours, about 0.5-1.5 hours or about 1-2 hours.

30. 30. The method of any one of claims 25 to 29, wherein the metathesis reaction conditions comprise converting potassium vanadate to the corresponding ammonium metavanadate compound and potassium salt.

31. 31. The method of any one of claims 25 to 30, wherein the metathesis reaction conditions comprise the sequential steps of adjusting the pH of the aqueous mixture to a range of about 8 to about 9, adding the ammonium salt to the aqueous mixture, and adding an ammonium Group VB metal compound species to the aqueous mixture at a pH in the range of about 7.5 to 8.5, preferably about 8.

32. 32. The method of any one of claims 25 to 31, wherein the crystallization conditions of the ammonium Group VB metal compound comprise a temperature in the range of from above 0°C to about 15°C or from above 0°C to about 10°C, a vacuum condition, and a crystallization period of from about 1 hour to about 6 hours, from about 1 hour to about 4 hours, or from about 1 hour to about 3 hours.

33. 33. A method according to any one of claims 25 to 32, wherein the conditions for filtering and washing the crystallised ammonium group VB metal compound comprise a wash temperature in the range of from above 0°C to about 15°C or from above 0°C to about 10°C, or a wash liquid temperature of about 10°C, preferably the crystallised ammonium group VB metal compound and the wash liquid comprise ammonium metavanadate, and optionally recycling the wash liquid for crystallisation of the ammonium group VB metal compound.

34. 34. The method of any one of claims 25 to 33, wherein the conditions for heating the ammonium Group VB metal compound comprise heating the ammonium Group VB metal compound at a temperature in the range of about 200-450°C, 300-450°C, 350-425°C, or about 375-425°C for a time sufficient to release ammonia in an amount of at least about 90%, 95%, 98%, or 99% of the amount present in the ammonium Group VB metal compound.

35. 35. The method of any one of claims 25 to 34, wherein the conditions for contacting the ammonium Group VIB metal compound filtrate with an inorganic acid comprise introducing the inorganic acid at a temperature in the range of about 50-80°C, 50-70°C or 55-70°C to achieve a pH of about 1-3, about 1-2 or about 1, and a reaction time of about 1 hour to about 4 hours, about 1 hour to about 3 hours, or about 1 hour to about 2 hours, preferably wherein the inorganic acid comprises or is nitric acid or sulfuric acid.

36. 36. The method of any one of claims 25 to 35, wherein the conditions for filtering and washing the Group VIB metal oxide compound precipitate with an ammonium Group VIB metal oxide compound washing solution comprise a washing temperature in the range of from above 0°C to about 25°C or from above 0°C to about 10°C, or a washing solution temperature of about 10°C, preferably the washing solution comprises Mo-depleted ammonium heptamolybdate at pH 1, and optionally recycling the washing solution for filtering and washing the Group VIB metal oxide compound.

37. 37. The method of any one of claims 25 to 36, wherein the recovery of Group VB metal present in the filtrate containing the Group VIB metal compound and the Group VB metal compound is greater than about 85 wt%, greater than about 90 wt%, greater than about 95 wt%, greater than about 97 wt%, greater than about 98 wt%, or greater than about 99 wt%.

38. 38. The method of any one of claims 25 to 37, wherein the recovery of the Group VIB metal present in the Group VIB metal compound and the filtrate containing the Group VB metal compound is greater than about 85%, greater than about 90% by weight, greater than about 95% by weight, greater than about 97% by weight, greater than about 98% by weight or greater than about 99% by weight.

39. 39. The method of any one of claims 25 to 38, wherein the saturated ammonium Group VB metal compound wash solution comprises the same ammonium Group VB metal compound as the crystallized ammonium Group VB metal compound, or the saturated ammonium Group VB metal compound of the wash solution is the same ammonium Group VB metal compound as the crystallized ammonium Group VB metal compound.

40. 40. The method of any one of claims 25 to 39, wherein the ammonium Group VIB metal oxide compound cleaning solution comprises the same ammonium Group VIB metal oxide compound as the deposited ammonium Group VIB metal oxide compound, or the ammonium Group VIB metal oxide compound of the cleaning solution is the same ammonium Group VIB metal oxide compound as the deposited ammonium Group VIB metal compound.

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