Eco-Friendly and Complex Processing of Vanadium-Bearing Waste for Effective Extraction of Valuable Metals and Other By-Products: A Critical Review

Abstract

Achieving the New World Sustainability Vision 2030 leads to enacting environmental restrictions, which aim to partially or totally reduce the negative impacts of different forms of waste and develop alternative technologies for eco-friendly and cost-effective utilization. Solid waste is a hazardous waste with many environmental and economic problems resulting from its storage and disposal. However, at the same time, these wastes contain many valuable elements. One of these solid wastes is heavy oil fly ash “HOFA” generated in power stations using heavy oil as fuel. HOFA is produced annually in massive amounts worldwide, the storage of which leads to the contamination of water resources by the contained heavy metals, resulting in many cancerogenic diseases. At the same time, these ashes contain many valuable metals in significant amounts, such as vanadium “V” and nickel “Ni” that can be extracted effectively compared to their low content and difficulty processing in their main ores. Hence, recycling these types of wastes reduces the environmental adverse effects of their storage and the harmful elements in their composition. This paper critically reviews the world resources of vanadium-bearing waste and various approaches described in the literature for recovering V, Ni, as well as other valuable metals from (HOFA) and other wastes, including pyro- and hydro-metallurgical processes or a combination. Hydro-metallurgical processes include alkaline or acidic leaching using different reagents followed by chemical precipitation, solvent extraction, and ion exchange to extract individual elements. The pyro-metallurgical processes involve the non-salt or salt roasting processes followed by acidic or alkaline leaching processes. The operational parameters and their impact on the efficiency of recovery are also discussed. The digestion mixtures of strong mineral acids used to dissolve metal ions in HOFA are also investigated. Bioleaching is a promising eco-friendly technology for recovering V and Ni through appropriate bacteria and fungi. Oxidation leaching is also a promising environmentally friendly approach and more effective. Among all these processes, the salt roasting treatment showed promising results concerning the cost, technological, and environmental effectiveness. The possibility of complex processing of HOFA has also been investigated, proposing innovative technology for completely utilizing this waste without any remaining residue. Effective zeolite for wastewater treatment has been formulated as a good alternative for conserving the available water resources.

1. Introduction

Vanadium is regarded as a rare metal, and its swift and consistently growing demand is challenged by its low concentrations in natural minerals. Vanadium extraction depends on the type of feed materials, efficiency, and technologies applied in processing raw materials. Therefore, recovering V and other valuable metals from secondary industrial resources like steel slag and ash containing these metal ions is highly intriguing regarding ecological and financial aspects. The increased industrial applications of vanadium contribute to the rising demand for vanadium, which has driven up the price of this metal. Currently, the value of vanadium ranges between $20,000 and $40,000 per metric ton [1,2]. These industrial uses encompass vanadium redox flow batteries, viewed as a promising solution to the global issue of large-scale energy storage production. They also include applications in the automotive and construction sectors, corrosion inhibitors, fertilizers, petrochemicals, catalysts for gas processing and sulfuric acid production, ferrovanadium alloys, thermistors, and more [3,4,5].

Unlike copper, nickel, or zinc, vanadium does not form concentrated deposits. Vanadium is found in many minerals, replacing Fe³⁺ or Al³⁺ in the octahedral positions of crystal lattices. This is possible because the ionic radii of V³⁺ and Fe³⁺ are almost the same, which means that vanadium can be found all over the Earth’s crust, accounting for approximately 0.014% [6]. Vanadium is found in nature in the form of oxides, sulfides, and phosphates, and it is often associated with other metals such as iron, titanium, aluminum, zinc, lead, and uranium. These compounds include vanadinite, titanomagnetite, descloizite, patronite, and carnotite. The concentrations generally have a V₂O₅ grade between 0.2 and 1%, while higher-grade deposits may have V₂O₅ levels exceeding 2% [7,8,9].

It is estimated that there are 26 million tons of vanadium reserves worldwide. The only vanadium mine-operating countries with the largest reserves are China with 9.5 million tons, Russia with 5 million tons, Australia with 6 million tons, and South Africa with 3.5 million tons. In contrast, the world’s resources total over 63 million tons [10]. Nearly all of the world’s vanadium production in 2022 came from China (68%), Russia (17%), South Africa (9%), and Brazil (6%). The total global production of vanadium this year was approximately 120 kilotons/year [8,10]. Over 88% of the total vanadium is primarily produced as co-products from slag during the smelting of titanomagnetite ores or steel refining, representing over 69% of the starting raw material and making them an essential source of vanadium. In contrast, approximately 19% of the vanadium was produced from primary vanadium ores and 12% from secondary sources (Figure 1) [1,8].

The world’s vanadium consumption reached 102.1 kilotons, representing an increase of nearly 45% over 2011. By the end of 2023, consumption is predicted to increase by up to 120~125 kilotons at a 5% annual growth rate, with increased demand for high-strength steel serving as the primary driver of this growth [8].

Coulsonite, a mineral of FeV₂O₄, is commonly found alongside magnetite and forms a series with FeCr₂O₄, also known as chromite [11]. The mineral ilmenite can also host vanadium, possessing a minor quantity [12]. Likewise, clay containing vanadium in weathered mineral deposits is commonly found, with V³⁺ taking the place of Al³⁺ in phyllosilicates and Fe³⁺ in goethite, FeOOH. A variety of vanadium oxides have been discovered in weathered zones of vanadium titanomagnetite deposits, including chlorite, (Fe, Mg, Al)₆(Si, Al)₄O₁₀(OH)₈, and goethite. It is worth noting that a lot of uranium deposits also contain vanadium in the form of minerals like carnotite (K₂(UO₂)₂V₂O₈.3H₂O) and tyuyamunite (Ca(UO₂)₂V₂O₈.5–8H₂O) that have uranyl vanadate [13].

From ores rich in minerals like carnotite (K₂(UO₂)₂V₂O₈·3H₂O), vanadium has been generated as a by-product of uranium mining at numerous mines in the southwestern regions of the USA [14]. After a moderate leaching process, a significant portion of the uranium deposits was dissolved. The remaining vanadium in the uranium-leaching residue was then extracted using a more aggressive leaching condition [1]. The processing of South Korean shale-hosted uranium/vanadium ore has been suggested using a similar strategy [15]. Liu [16] has also been found to have comparable deposits in Western Australia. Coal and bituminous shale also contain trace levels of vanadium [17]. In northern Queensland, an abundant oil shale deposit at Julia Creek has been found to include vanadium and trace quantities of molybdenum [18].

Trace amounts of Vanadium have also been found in crude oil [19,20,21,22], among other potentially essential sources. Vanadium levels in crude oil are very high, reaching up to 0.05% in certain regions of Venezuelan oil [23,24]. During oil combustion, its vanadium content is transferred to the ash, which contains up to 10~18% of V₂O₅, making it an attractive material for vanadium recovery [25,26,27].

According to a report by White and Levy [28], the main sources of vanadium released into the environment are steel co-production, which accounts for 75% of the total, crude oil processing, which accounts for 10%, and mining, which accounts for less than 10%. In the Middle East and Africa, fuel oil and crude oil are easily accessible resources for power plants and desalination plants. For instance, it has been approximated that Saudi Arabia, with the world’s largest oil reserves, consumes more than 40 million metric tons of crude oil and heavy fuel oil annually for thermal power plants and seawater desalination [29]. Egypt, Africa’s largest non-OPEC oil producer, produces approximately 11 million metric tons of heavy fuel oil (HFO) annually, of which approximately 9 million metric tons are consumed domestically [30,31,32]. Petroleum and other liquids contribute 36%, and natural gas 57%; they were considered the most-consumed fuels in Egypt in 2020, according to BP’s Statistical Review of World Energy 2021, in addition to renewable energy and coal, which accounted for 6% and 1% of the country’s total energy consumption [33,34]. The country implemented measures to decrease its natural gas usage in power generation, which typically accounts for up to 90% of its total generation. As part of this effort, the country has significantly increased its use of oil-fired generation, resulting in an average fuel oil burn of over 100 kb/d in the first nine months of 2022, the highest level seen in four years [29].

To date, in many developing countries, most of the generated ashes do not have any industrial application but are only stored in landfills. Compared to other V resources, which are relatively limited and low-grade, these materials have no mining or operating costs. The disposal of ash or steel slag is a major environmental problem facing thermal power plants and iron production industries, and no monitoring is taken to prevent the contamination of groundwater and soil [30,35]. It is recommended to continuously monitor the metals of environmental impact, such as nickel, vanadium, and zinc, in ash samples due to carcinogenic and health hazards. Therefore, it is recommended to consistently monitor and recover valuable metals from industrial waste, as this is highly beneficial for both environmental protection and economic considerations. This review aims to offer a summary of advanced technologies used for vanadium recovery from various types of vanadium-bearing waste, as well as the comprehensive utilization of their solid residue.

2. Chemical and Mineralogical Composition of Vanadium-Bearing Waste

Table 1. Chemical composition of vanadium-bearing slag generated from various industrial plants in different locations in the world.

Sources	Technology	Chemical Composition					Ref.
		V2O5/V2O3	FeO/Fe2O3	Cr2O3	TiO2	MnO/MnO2
Panzhihua Iron and Steel	BF-BOF	8.5–14.3	32.9–42.1	2.1–4.3	11.1–12.9	5.9–9.1	[36]
Kapok Iron and Steel	BF-BOF	14.2–16.0	31.9–32.3	-	12.0	7.6–8.7	[37]
Chengde Xinxin Vanadium and Titanium Chemical	BF-BOF	8.5	51.1	3.5	10.5	5.2	[38]
Desheng Iron and Steel Group	BF-BOF	15.2–20.8	35.3–64.5	7.7–10.6	6.39–8.5	5.5–7.6	[39]
Chengde Iron and Steel	BF-BOF	10.2–13.4	36.7–49.1	1.7–4.2	6.8–11.1	5.2–7.2	[40]
Esfahan Steel	BF-BOF	1.5–1.9	17.3–17.6	-	1.0–1.5	4.2–4.5	[41]
Sichuan Weiyuan Iron and Steel	BF-BOF	14.3	24.8	4.4	7.4	8.5	[42]
Pan Steel	DR-EAF	8.9	24.3–25.2	2.0–8.7	3.3–14.7	1.6–13.8	[43]
CITIC Jinzhou Ferroalloy	DR-EAF	15.3	30.5	2.3	13.7	10.9	[44]
British Steel	BF-BOF	0.82	32.0	0.2	0.3	4.5	[45]

BF = blast furnace; BOF = basic oxygen furnace; DR = direct reduction; EAF; electric arc furnace.

shows the chemical composition of vanadium-bearing waste produced from different iron and steel production facilities in different locations worldwide. The chemical structure of vanadium oxides, such as V₂O₃ or V₂O₅, varies based on the type of feed materials introduced into the furnaces, the efficiency of the processing methods, and the desired final products. Typically, CaO, SiO₂, MgO, and Al₂O₃ contained in the vanadium-bearing slag are in the ranges 2~45%, 6~25%, 1~16%, and 2~10%, respectively. During reduction smelting, chromium and titanium oxide are also found in the vanadium-bearing slag, so obtaining pure vanadium requires critical separation of those metals after extraction.

The chemical structure of HOA reveals that it primarily consists of carbon, along with several metallic elements.

Table 2. Chemical composition of vanadium-bearing waste reported in different studies.

	References
Element (ppm)	Ref. [53], PP	Ref. [53], DP	Ref. [21]	Ref. [54]	Ref. [55]	Ref. [56]	Ref. [57]	Ref. [58]	Ref. [35]	Ref. [59]	Ref. [25]
Carbon (%)	90.40	56.70	85.56	51.86	-	67.4	-	77.40	-	14.73	18.65
Sulfur (%)	7.77	27.18	-	-	-	8.6	-	7.10	-	5.83	11.23
Vanadium	9072	31,044	2958	34,487	50,000	38,000	7670	12,900	3540	85,000	49,100
Nickel	2382	13,633	1762	11,852	15,400	16,000	18,000	6800	1055	52,500	24,200
Cadmium	1.65	3.7	3.28	1.59	-	-	-	-	60.4	-	920
Arsenic	2.54	1846	2.24	68.29	-	-	-	-	128.5	-	1.61
Cobalt	2.88	12.33	3.28	247.79	2200	-	-	-	60.7	-	2140
Chromium	36.79	113.09	4.06	107.60	8000	-	-	-	70.4	-	-
Selenium	1.00	6.81	11.60	13.20	-	-	-	-	7.25	-	-
Lead	17.09	13.94	11.00	116.10	-	-	-	-	27.85	7200	1900
Zinc	21.92	118	130	592.10	-	-	1110	4000	65.5	38,000	7600
Copper	10.44	50.40	170.4	120.30	-	-		17,000	57.12	290	1760
Iron	7210	8771	-	-	220,500	8000	59,800	1400	176.5	97,400	59,600
Magnesium	6971	94,608	-	-	2000	-	3430	14,100	3615	5920	3.6
Manganese	23.9	149.26	-	-	-	-	-		12.8	900	2.4
Calcium	582.3	4121.2	-	-	-	-	-	2300	3380	2100	2354
Sodium	1395	7555	-	-	-	19,000	-	-	1310	2045	1524
Aluminum	3541	1041.8	-	-	-	1040	2870	2500	642.4	7123	4252
Barium	7.42	49.68	-	-	-	-	-	-	69.0	-	5214
Others (ppm)	-	-	Hg, 0.25	-	Mo, 3500	Si, 8000	-	Si, 800; O, 93 200	-	Si, 39,000	Si, 36,800

PP = power plant; DP = desalination plant.

presents the chemical composition of ash samples generated from burning heavy oil in various power and desalination plants across different global locations. Industrial ashes primarily consist of vanadium (V) and nickel (Ni), along with toxic elements like chromium, arsenic, cadmium, lead, cobalt, and selenium. The maximum content of V and Ni in ash is 5 wt.% and 1.8 wt.%, respectively [35]. Heavy oil ash chemical composition also differs based on the makeup of the heavy fuel oil, parameters of operation, and chemicals used [35,46]. Two types of ash (boiler ash and fly ash) are generated by the combustion of HFO in power plants, which is expected to increase as energy consumption rises in the next few years. Around 3 kg of ash residue is produced when 1000 L of heavy oil is burned [47]. A higher level of V and Ni was found in heavy oil ash (HOA) produced by various power plants worldwide [48]. The ash fraction is too heavy to be introduced into the flue gas and hence deposited outside the boiler tube, which is known as the boiler ash. It is usually classified as high-grade ash which contains (4.4~19.2%) V, (2.7~8.5%) Ni, and burned C [49,50]. The other type of ash is known as fly ash, which is fine combusted residue usually collected from the stack or electrostatic precipitators (ESPs). It is usually classified as low-grade ash which contains (0.3~1.5%) V, (0.2~0.5%) Ni, and unburned (30~80%) C [51]. The refining of crude oil often entails some kind of thermal cracking to produce a light oil fraction and a heavier residual oil. Most metals and a small amount of sulfur are still linked to the non-volatile elements and end up in a residual oil fraction. Several factors influence the elemental composition of heavy fuel oil, including the source of crude oil, the yield of HFO from crude oil, and the degree of processing [52].

Table 3. Characterization of the minerals in vanadium-containing waste produced from various industrial plants in different locations in the world.

Source	Mineralogical Phases	Ref.
Panzhihua Iron and Steel	(Fe,Mn)V2O4,(Fe,Mn)2SiO4,CaMn(SiO3)2	[36]
Kapok Iron and Steel	(Fe, Mn)O, (V,Ti)2O3, 2(Fe,Mn)O, SiO2, SiO2	[37]
Chengde Xinxin Vanadium and Titanium Chemical	(Mn0.84Fe0.16)(Mn0.16Fe1.34Cr0.5)O4, Fe2.1Ti0.74Mn0.02V0.01Ca0.01Si0.01Al0.05Mg0.06O4,Fe2SiO4, SiO2	[38]
Desheng Iron and Steel Group	(Mg,Fe)(V,Cr)2O4, Fe2SiO4, Ca(Fe,Mg)Si2O6,Fe2TiO4	[39]
Chengde Iron and Steel	(Mn,Fe)(V,Cr)2O4, Fe2SiO4, SiO2	[40]
Esfahan Steel	Ca(OH)2,Ca3SiO5, CaFeO2,Ca2V2O7	[41]
Sichuan Weiyuan Iron and Steel	(Mn,Fe)(V,Cr)2O4,Fe2SiO4, (Fe,Mg)Si2O6,Mg2TiO4	[42]
Pan Steel	(Mn,Fe)(V,Cr)2O4,Fe2TiO4, (Fe,Mn)2SiO4	[43]
CITIC Jinzhou Ferroalloy	(Mn,Fe)(V,Cr)2O4,CaFeAlSiO6,Fe2.5Ti0.5O4	[44]
British Steel	MgO, Al2O3, β-Ca2SiO4, Ca2(Al,Fe)2O5, FeO, Mg(OH)2, C	[45]
Fiume Santo-Porto Torres	MgSO4.H2O, (VO2)2H2(SO4)3, VOSO4.nH2O CaSO4, NiSO4.nH2O	[64]
El Kriymat Electric power station	(VO2)2, FeV2O4, Mg2SiO4, Ca2V2O7, (Ca(Fe,Al)2SiO6), NiSO4.H2O6	[22,25]

presents the mineralogical compositions of vanadium-bearing slag produced by different iron and steel-making facilities and ashes. The spinel phases of [(Fe, Mn)₂(V, Ti) O₄] and [(Fe, Mg, Mn) (V, Cr)₂ O₄] contain vanadium, titanium, and chromium. The spinels are distributed throughout the olivine matrix (Fe, Mg)₂SiO₄, augite (Ca(Fe, Mg)Si₂O₆), fayalite (Fe₂SiO₄), and diopside phases (Ca(Fe, Al)₂SiO₆) [60,61,62]. Previously reported data indicated that V-Ti and V-Cr spinels are arranged in a layered core configuration. Chromium is primarily located in the central core of the grains, vanadium is prevalent in the outer layer, and titanium is situated along the spinel grain boundary. This distribution reflects the varying crystallization abilities of each spinel, ranked as Cr > V > Ti [43]. Other elements, such as manganese, are evenly dispersed within the spinel grain. Magnesium is concentrated alongside chromium, forming the minor phase of MgCr₂O₄ [60]. The Ca₂V₂O₇ phase of vanadium was also documented [41], which resulted from the substantial amount of limestone utilized in the iron and steel production process [63].

Additional oxides, originating from the fluxes added to the furnace charge, have also been identified. Dondi [64] states that X-ray diffraction patterns show that two types of fly ash (ash B from Apulia and ash F from Sardinia) produced from the combustion of orimulsion in Italian thermal power plants have a complex composition. These fly ashes contain high grades of V, Ni, S, and Mg and exhibit increased reflections, several line interferences, and overlaps. This complexity makes it difficult to make certain attributes. The interpretation of these patterns indicates that magnesium sulfate monohydrate (MgSO₄.H₂O) and hydrogen vanadyl sulfate [(VO)₂H₂(SO₄)₃] are predominated, along with magnesium monoxide (MgO), anhydrous, monohydrate, and pentahydrate vanadyl sulfate (a-VOSO₄, VOSO₄.H₂O, VOSO₄.5H₂O). These phases are associated with minor amounts of calcium sulfate (CaSO₄) and nickel monoxide (NiO), and presumably sodium aluminum silicate (NaAlSi₃O₈) and traces of magnesium sulfate hexahydrate (MgSO₄.6H₂O). Other analyses of the insoluble residues indicated that in addition to magnesium, nickel oxides, and anhydrous vanadyl sulfate, other vanadium compounds, i.e., VO₂, ꞵ-V₂O₅, VOOH, and VO(OH)₂, were present in small amounts of untreated ash, which cannot be detected in the collective sample due to their low concentration. However, identifying these compounds suggests the presence of V³⁺ and V⁵⁺, together with the monopolized vanadyl ion valence of V⁴⁺. The existence of these vanadyl ions can vary depending on the oxidation conditions during the combustion of heavy oil or the cooling techniques used [65,66].

3. Vanadium Processing Technologies

As previously mentioned, steel slag and oil ash are important secondary sources of valuable elements (V, Ni, Cr, and Zn). However, there are two main ways for processing industrial solid waste to extract vanadium and other useful metals. The first method, direct leaching, involves using acid (such as H₂SO₄, HCl, and HNO₃) or alkaline (such as NaOH and Na₂CO₃) solutions to dissolve the valuable metal from vanadium-bearing waste. The dissolved metals are then recovered from the solution using various techniques such as chemical precipitation, solvent extraction, or ion exchange. The second is the roasting-assisted leaching process, including salt-free, calcium, or salt roasting, in which the solid waste is heated to a high temperature to allow the separation of valuable vanadate materials. Acid, alkaline, and water are frequently employed as leaching agents, as outlined below [67,68,69,70,71,72,73,74]. A comprehensive flowchart for recovering V from solid waste is shown in Figure 2 [35]. After extracting vanadium and nickel as their respective chemical compounds (NH₄VO₃ and NiC₂O₄), calcination is conducted to produce V₂O₅ and NiO.

3.1. Direct Leaching Process

Direct acidic or alkaline leaching is used to dissolve almost all valuable metals into the solution from various types of solid waste and then extract these metals separately using adequate techniques [75,76,77,78,79,80,81,82].

3.1.1. Acid Leaching

H₂SO₄ is a cost-effective option for leaching procedures, which makes it a popular choice due to its economic benefits. The H₂SO₄ acid has the ability to dissolve several metallic cations like V, Ni, Fe, and Mg, among others, creating metal sulfates. The leaching of significant economic and nuclear elements such as uranium, hafnium, and zirconium, as well as vanadium and nickel, with or without oxidant agents (NaClO, MnO₂, or KClO₃), has been investigated by Abd El-Hamid [83]. It was noted that vanadium leaching efficiency of 95% at a concentration of 200 g/L H₂SO₄, 6% (solid/solid) manganese dioxide, 80 °C, 6 h, 1/10 solid/liquid ratio of −200 mesh ore granule size. According to Deng [84], the direct acid leaching technique could increase the vanadium leaching efficiency from 14 to 73% by adding an oxidant. Barik [85] created a technique for extracting vanadium (V) and nickel (Ni) from solid industrial wastes abundant in these metals. The procedure includes leaching with sulfuric acid (H₂SO₄), followed by solvent extraction, precipitation, and crystallization. It was noted that using 1.35 M H₂SO₄ at 40 °C for 90 min, both V and Ni were extracted with a 98% recovery rate. Vanadium is selectively extracted from the pregnant solution using 40% LIX 84-I at pH 0.5, and then by adding an aqueous solution of ammonia NH₃; it precipitates as ammonium metavanadate (NH₄VO₃). Ammonium oxalate (NH₄)₂C₂O₄ is used to selectively separate Ni that has been converted into Ni oxalate in the raffinate (NiC₂O₄). Finally, NH₄VO₃ can be heated to produce vanadium pentoxide (V₂O₅) [86]. On the other hand, nickel oxide is obtained by heating nickel oxalate at 450 °C for 2 h.

Nazari [87] optimized the parameters of the H₂SO₄ leaching process to maximize V and Ni recovery from fly ash samples. The maximum recovery of 94% V and 81% Ni has been achieved at the following optimal conditions: 19.5% H₂SO₄ concentration at 80 °C, 9.15% S/L ratio by weight, and for 2 h. Fly ash was subjected to a three-step procedure, which included leaching with H₂SO₄, oxidative precipitation, and washing to extract V as V₂O₅ from the fly ash, and about 90% of V was successfully recovered. Boiling 1 M H₂SO₄ for 30 min. at 3 mL/g L/S ratio is considered to be the operational parameter for the leaching process. During the oxidative precipitation process, sodium chlorate (NaClO₃) was added as an oxidizing agent of V⁴⁺ to V⁵⁺. The pH dropped because of this oxidation of V, which also released H into the environment; hence, the appropriate pH was maintained by adding Na₂CO₃. In order to remove Na and other impurities, the V₂O₅ precipitate was then washed with an acidic solution [26]. Leaching experiments on HOA were performed by Navarro [88] in 0.5 M H₂SO₄ solution at 4 mL/g as L/S ratio at 200 rpm agitation speed for 24 h and at room temperature. After the leaching procedure, the material was washed for 1 h. in water at 4 mL/g L/S ratio. Ni recovery was only about 12% under these conditions, but V reached 98%.

Aburizaiza [89] detailed a multi-step approach for leaching and extracting V, Ni, and Fe from HOA, which involves the following steps: (i) Acidic leaching of fly ash in a solution containing mineral acids like perchloric acid (HClO₄), hydrochloric acid (HCl), and hydrogen peroxide (H₂O₂); (ii) Ammonium pyrrollidine dithiocarbamate (C₅H₉NS₂.NH₃) in chloroform (CHCl₃) is combined and agitated with the leachate during the organic extraction of the metals. Once equilibrium is reached, a layer of organic phase that contains the metal cations is formed. The findings demonstrate that the organic phase successfully extracts the targeted metal ions (V, Ni, and Fe); (iii) the V, Ni, and Fe metal ions are removed from the organic phase by combining them with 1 M HNO₃ that contains Hg²⁺ ions. Therefore, metallic cations (V³⁺, Ni²⁺, and Fe²⁺) are present in the final aqueous nitric acid solution; (iv) The addition of dimethylglyoxime (C₄H₈N₂O₂) in chloroform allows the extraction of the total Ni from the stripped nitric acid solution while leaving the V³⁺ and Fe²⁺ ions dissolved in the aqueous nitric acid solution phase. In order to recover nickel in the form of NiO, the Ni-containing organic phase is separated and evaporated till it is dried; (v) Total Fe is extracted from the aqueous nitric acid solution phase produced in step (iv) by 4-pypyridyl treatment. The organic phase that contained Fe was separated, evaporated, and ignited to produce Fe in the form of FeO; (vi) vanadium as V³⁺ and other trace elements can be found in the aqueous solution that was left over. The study does not present an efficient technique for extracting V from the final aqueous solution. Additionally, this approach demands significant amounts of chemicals to handle a small quantity of fly ash.

Tokuyama [58] studied the dissolving of HOA in two steps. In the first step, Ni, Zn, Mg, Al, and Fe are dissolved in water, and in the second stage, V is dissolved in H₂SO₄, as shown in Figure 3. Moreover, it was found that HCl and H₂SO₄ have similar leaching efficacy. In the same study, leaching with concentrated NaOH might result in the selective leaching of V.

Tsygankova [48] concluded that the HOA leaching process depends on both its phase change and chemical composition. The outcomes of acidic leaching for three distinct types of fly ash, each with different chemical and phase compositions, have been examined. Each type of fly ash requires specific leaching conditions, involving 5 to 9% H₂SO₄ concentration, with temperatures set between 20 and 80 °C, and durations spanning from 30 to 60 min. Vanadium is precipitated as V₂O₅ by oxidizing the leaching solution with H₂O₂ at 20 °C for 30 min. After further heating to 95 °C and at a pH range from 1.8 to 2, precipitation is accomplished.

Khalafalla [90] examined the extraction of V, Ni, and Zn by leaching the heavy oil ash (HOA) in H₂SO₄ solution, as shown in Figure 4. High-tension Magnetic Separation was used to physically concentrate the fly ash, increasing the amounts of V₂O₅, ZnO, and NiO, to 18.1%, 8.11%, and 11.18%, respectively. Following that, leaching was performed using a 180 g/L H₂SO₄ solution with 4% MnO₂ as an oxidant for 10 h at 80 °C which resulted in the recovery of 96.5% of vanadium, 94.8% of nickel, and 99.1% of zinc. Ultimately, vanadium was extracted from the leach solution using 3% Alamine 336 in kerosene as the solvent, while nickel and zinc remain dissolved in the raffinate. After vanadium was extracted from the organic phase, V₂O₅.H₂O was precipitated by adding H₂SO₄. In contrast, when Na₂S was added to the raffinate solution, Ni and Zn co-precipitated, forming a (Zn-Ni) sulfide cake. This cake could be suggested for hydro-metallurgical techniques to recover and separate Zn and Ni [91].

Rahimi [92] developed a distinctive technique for extracting vanadium from HOA utilizing lemon juice and organic acids. The citric acid (90 mg/g), malic acid (0.86 mg/g), and ascorbic acid (1.24 mg/g) are all present in lemon juice. The idea was to obtain organic acids by an ecologically acceptable method as an alternative to bioleaching, which produces such acids through extremely sluggish microbial development. The recommended conditions for the maximum recovery of 88.7% V were 2 h. ultrasonic leaching of HOA in a 27.9% lemon juice solution with 10% hydrogen peroxide at 35 °C and 0.01% S/L ratio. Sodium carbonate (Na₂CO₃) was added to the liquid to raise the pH level to 9 or 10, and the addition of CaCl₂ precipitates Al and Fe. Next, ammonium chloride (NH₄Cl) was added, leading to the precipitation of V as pure NH₄VO₃, then calcined at 500 °C to obtain vanadium pentoxide (V₂O₅). Eventually, the results revealed that assisting leaching agents, such as H₂O₂ and ultrasonic, is essential in accelerating V dissolution’s kinetics.

High-pressure and high-temperature leaching demonstrates significant improvements compared to traditional acid leaching. For instance, according to Mu [93], 97.9% vanadium and 94.5% iron were leached out at a temperature of 140 °C, a concentration of 200 g/L H₂SO₄, a partial pressure of 0.5 MPa of oxygen, a period of 120 min, and an L/S ratio of 20:1. These are in completely agree with the findings published by Zhang [94,95,96]. Zhou [97] studied the dissolution kinetics of vanadium trioxide in sulfuric acid-oxygen medium. The highest dissolution rate was obtained with increased stirring speed before 800 rpm, oxygen partial pressure in the range of 0.6~1.4 MPa, and particle size decreased. Still, it was nearly independent of the sulphuric acid concentration from 0.4~2.0 mol/L and stirring speed over 800 rpm. Amer [98] recovered V₂(SO₄)₃ and NiSO₄ from Egyptian boiler ash by direct leaching using H₂SO₄ under oxygen pressure. The recommended leaching conditions have been determined to be 200 °C, 15 min, 15 bar of oxygen pressure, 60 g/L of H₂SO₄ concentration, and 1 S/L ratio. Adding an aqueous solution of ammonia NH₃ and neutralizing the acidic sulfate solution, V is precipitated as vanadium hydroxide (V(OH)₃). On the other hand, metallic Ni can be electrodeposited from the NiSO₄ electrolyte. The results of comprehensive studies employing acidic leaching methods on various vanadium resources are shown in

Table 4. Results of different studies on direct acidic leaching of vanadium from various V-bearing waste.

Ref.	Acid Leaching					Leaching Efficiency
	Lixiviant	Conditions	Temp. °C	Time	S/L Ratio
[98]	H2SO4	−250 µm, 60 g/l H2SO4, 15 bar of O2 pressure	200 °C	15 min	1:1	V = 95%
[85]	H2SO4	1.35 M H2SO4, 300 rpm stirring	40 °C	90 min	1:5	V = 98% Ni = 98%
[87]	H2SO4	−75 µm, 19.5% H2SO4 Conc.	80 °C	120 min	1:9	V = 94% Ni = 90%
[26]	H2SO4	1 M H2SO4, 300 rpm stirring	100 °C	30 min	1:3	V = 90%
[88]	H2SO4	−500 µm, 110 °C drying, 0.5 M H2SO4	Room Temp.	24 h	1:4	V = 98% Ni = 12%
[89]	H2SO4 + HCl + HNO3	1.0 M Conc., 300 rpm stirring	RT	6 h	1:5	V = 98% Ni = 95% Fe = 86%
[48]	H2SO4	−75 µm, 5~9% H2SO4 Conc.	20~80 °C	30~60 min	1:4	V = 96%
[90]	H2SO4	180 g/L H2SO4, 4% MnO2, 500 rpm stirring	80 °C	600 min	1:10	V = 96.5% Ni = 94.8%
[92]	Ultrasound, H2SO4	−75 µm, 27.9% lemon juice, 10% H2O2	35 °C	120 min	0.01%	V = 88.7%
[99]	H2SO4	−100 µm, 0.5 N H2SO4, 400 rpm stirring	30 °C	120 min	1:5	V = 65% Ni = 60% Fe = 42%

.

3.1.2. Alkaline Leaching

Several investigations on direct alkaline leaching of V-bearing waste have been carried out, indicating that Ni is insoluble in alkali solutions, which are more suitable for selective leaching of V to increase recovery. A two-stage leaching approach for HOA was proposed by Tsai [99]. The fly ash is leached in ammonia water containing ammonium sulfate (NH₄)₂SO₄, and the Ni is extracted from the solution in the first stage. The second step involves recovering V by leaching the ash residue in an alkaline (NaOH) solution. However, the method was not used in more recent research because of the decreased recovery of Ni (60%) in the ammonia/(NH₄)₂SO₄ solution, which also dissolves 8% of V in the first step. Furthermore, the study did not propose or conduct a method to recover V and Ni from the leaching solution. In a power plant, alkaline leaching and sodium hydroxide “NaOH” solution were used by Al-Zuhairi [100] to extract V with a high recovery yield (98%) from red crude oil residues. According to their discussion, 2 M NaOH concentration at 100 °C for 2 h is the ideal extraction environment for V. The effluent is then acidified to lower the pH and treated with aqueous solution of ammonia (NH₃), to precipitate V as NH₄VO₃.

According to Navarro [88], alkaline leaching of HOA by NaOH is favored over acidic leaching because it is more selective for V, which is less efficient than acidic leaching by H₂SO₄. Alkaline leaching with NaOH is carried out using a 2 M NaOH solution at 1/4 g/mL S/L ratio for 24 h at room temperature and stirring at 200 rpm. Leaching is followed by three times for 1 h water rinses. At these conditions, vanadium recovery achieves 90%, but repeated leaching and washing 6 times, each for 1 h, enhances V recovery to 98%. When the contact time approaches 12 h, it is observed that the temperature has a minimal impact on the effectiveness and kinetics of V leaching. In order to improve selectivity for V recovery while preventing Si leaching, sodium carbonate (Na₂CO₃) is used for leaching HOA. However, by employing 0.66 M Na₂CO₃ solution in the leaching procedure under the same conditions as for NaOH leaching, 80% of V is recovered. Leaching with Na₂CO₃ exhibits more excellent vanadium selectivity while having a lower leaching efficiency than leaching with NaOH.

Hakimi [101] established a flowsheet for extracting V with a high percent recovery from V-rich HOA using NaOH solution at 95 °C for 4 h. as shown in Figure 5. Leaching is followed by adding 4.5 M H₂SO₄ solution to neutralize the alkaline leachate. Al and Si are deposited once the pH is raised to 8, leaving V dissolved as sodium vanadate in the filtrate. After being treated with an ammonium compound, the latter precipitates V as NH₄VO₃, which is then heated at 450 °C for 1 h to form V₂O₅.

Akita [102] studied the recovery of nickel and vanadium from fly ash using a two-step leaching process. The first stage involves leaching Ni with NH₄Cl solution, while the second step involves leaching the remaining ash with Na₂CO₃ to dissolve the vanadium. Subsequently, vanadium is extracted using a solvent mixture of TOA and toluene, and then precipitated as NH₄VO₃ using NH₄Cl. However, nickel can be extracted as NiS from its leach solution by precipitating it with Na₂S, as illustrated in Figure 6.

Vanadium, nickel, and molybdenum recovery from HOA was accomplished by Stas [55], using a two-stage leaching procedure, as shown in Figure 7. The first step is the alkaline leaching using NaOH solution to extract V and Mo. In the subsequent second step, acid leaching using H₂SO₄ leaches the residual fly ash to recover Ni. The maximum V recovery of 90% is obtained at 5 mL/g L/S ratio, at 100 °C, and for 3 h. The alkaline leaching solution containing V and Mo is gently agitated and cooled to 5 °C for one hour, which causes V to precipitate as sodium vanadate. This precipitate is then removed through filtration, leaving behind a filtrate that contains sodium molybdate. The later filtrate is heated to 90 °C and acidified with HNO₃ to precipitate Mo as H₂MoO₄. Conversely, sodium vanadate is dissolved again in a 5% HNO₃ solution (pH = 8), and (NH₄)₂SO₄ is introduced to the solution to precipitate vanadium as N₄HVO₃. The latter is converted to V₂O₅ after being calcined at 500 °C for 24 h. The remaining ash is leached using a 5 M H₂SO₄ solution for 3 h at 100 °C and a 4 mL/g L/S ratio, which leads to extraction of 80% in the second acidic leaching stage. The pH is subsequently increased by adding a NaOH solution to the acidic leachate in two stages, which results in the precipitation of Fe and the residual V. Finally, sodium carbonate is introduced to precipitate nickel in the form of nickel carbonate (NiCO₃).

Al-Ghouti [103] studied the recovery of vanadium and nickel from HOA by two leaching processes. Using NH₄Cl/NH₃ solution made up of 2 M NH₄CL and 2 M NH₃ at 50 °C, at 19 mL/g 1 L/S ratio, and for 6 h, Ni is selectively extracted in the first stage. After treating the leachate with sodium sulfide (Na₂S), nickel is precipitated as NiS with 56% recovery. In the following step, vanadium is extracted from the nickel-free remaining ash using agitation leaching in a 2 M Na₂CO₃ aqueous solution at 70 °C and a pH of 5.5 for a duration of 6 h. However, along with V, Fe, Mg, and Ca are also removed by Na₂CO₃ solution. Triethylamine/toluene solution [0.1 M of (CH₃CH₂)₃N in toluene] is added to the Na₂CO₃ solution to perform selective extraction of V. The organic layer is removed and then blended once again for 3 h with a fresh 2 M Na₂CO₃ solution to extract V from the aqueous layer. Finally, the resultant aqueous layer is mixed with 1.3 M NH₄Cl for 20 h to achieve maximum recovery of 45% by precipitating the extracted V. Alkaline pressure leaching was carried out by Shahnazi [104] with a maximum vanadium leaching of 90% under the concentration of 40 to 50% NaOH and a temperature range of 110 to 160 °C. In the same way, Qiu [105] examined how pure vanadium trioxide dissolves in an oxygen system under alkaline conditions. The findings indicated that higher temperatures (130 °C) and oxygen pressure (700 kPa) resulted in a faster rate of vanadium dissolution. Additionally, stirring speeds above 800 rpm, and sodium hydroxide concentrations at 1.0 M were found to maximize the dissolution rate of vanadium (90%). A novel approach utilizing NaOH-NaNO₃ binary melts has been suggested for treating vanadium slag [106]. This method enables the extraction of both vanadium and chromium during the leaching process. When the reaction conditions are optimized (1:1 NaOH-NaNO₃ solid ratio, 400 °C, 0.5 L/min atmospheric pressure for 2 h), vanadium and chromium recovery rates can reach 93.7% and 88.2%, respectively, within 6 h. The results indicate that the reaction temperature and KOH-to-ore mass ratio are more influential factors for the extraction of vanadium and chromium.

Table 5. Results of different studies on the alkaline leaching of vanadium from various V-bearing wastes.

Ref.	Alkaline Leaching					Leaching Efficiency
	Lixiviant	Condition	Temp. °C	Time	S/L Ratio
[99]	NaOH	−250 µm, 2 N NaOH, pH 14, 400 rpm agitation speed	30 °C	120 min	1:5	V = 80%
[99]	NH4OH	−250 µm, 4 N NH4OH, pH 10	30 °C	120 min	1:5	V = 50% Ni = 60%
	NH4OH + (NH4)2SO4 + NaOH	0.25 N NH4OH, 4 N (NH4)2SO4, pH 8.5, 2 M NaOH	30 °C	120 min	1:5	V = 80% Ni = 60%
[100]	NaOH	2 M H2SO4, 300 rpm stirring	100 °C	120 min	1:3	V = 98%
[88]	NaOH	2 M NaOH+ H2O, 200 rpm agitation speed	RT	24 h	1:4	V = 98%
	Na2CO3	0.66 M Na2CO3, 200 rpm stirring	RT	24 h	1:4	V = 80%
[101]	NaOH	−500 µm, 110 °C drying, 0.5 M H2SO4	100 °C	240 min	1:2.7	V = 99.6%
[102]	HN4Cl + NH4OH	2 M HN4Cl + NH4OH	50 °C	300 min	1:5	Ni = 59%
	Na2CO3	2 M Na2CO3	70 °C	240 min	1:4	V = 63%
[103]	NH4Cl + NH3	2 M NH4Cl+ 2M NH3	50 °C	300 min	1:19	Ni = 56%
	Na2CO3	2 M Na2CO3, 200 rpm stirring	70 °C	300 min	1:20	V = 45%
[55]	NaOH + H2SO4	8 M NaOH 5 M H2SO4,	100 °C 100 °C	180 min	1:5 1:4	V = 90% Ni = 80%

summarizes the results of in-depth investigations on different vanadium resources using the alkaline leaching technique.

Even though direct acid (H₂SO₄, HCl, and HNO₃) [107,108,109] and alkaline (NaOH, NH₄OH, NaCO₃, and NH₄Cl) [110,111,112] leaching methods are the cheapest operational and energy-efficient ways to extract valuable metal ions from heavy oil ash. They have several drawbacks, including a lack of selectivity in leaching the target metals, requiring large amounts of high acid concentration, using an oxidizing agent, polluting metals in solid ash residue, and producing leachate with hazardous elements that can complicate the precipitation and purification processes [4,113,114,115]. Due to the lack of an economically feasible alternative for processing vanadium-bearing waste, roasting is utilized to enhance the leaching process.

3.1.3. Bioleaching Process

Choosing the right bioleaching conditions is crucial for the microorganisms to grow ideally. The process has the advantage of being low in production cost and having a minimal environmental impact [116,117,118,119,120]. Bacteria like Acidithiobacillus, A. ferrooxidans, and Acidithiobacillus thiooxidans are well suited for chemolithoautotrophic applications in acidic conditions [121,122]. A study using A. ferrooxidans discovered that after 14 days, the roasted slag could leach out 83% of the chromium while only 20% of the vanadium was leached out [117]. Gomes’s [123] results align with those obtained from a modified acidophilic culture of A. thiooxidans and A. ferrooxidans. Under alkaline conditions, high pH-tolerant organisms are being assessed [116,117]. These include Pseudomonas putida, a heterotrophic bacterium, and Aspergillus niger, a heterotrophic fungus. A study involving P. putida revealed that after 15 days, 75% of vanadium was leached from a roasted slag sample. Using the bacterium on the calcine increased dissolution to 90% [107,117].

3.1.4. Electro-Oxidation Leaching

In the leaching reactor, electrodes are inserted, and an electric field is used to convert low-valent vanadium to a higher oxidation state. Outstanding outcomes were obtained, with high recoveries of 95% for vanadium and 90% for chromium. The alkaline leaching system was applied with current densities ranging from 750 to 1000 A/m², NaOH concentration of 40~50 wt.%, and 90~120 °C temperature [124,125,126]. Both macro and micro perspectives allow for an understanding of the electrooxidation system’s assistance in leaching. The schematic of both mechanisms is explained in more detail by Lee et al. [127]. Diffusion of reaction products is the focal point of the micro-viewpoints [124]. During leaching, metallic ions escape from the iron phase more readily in an electrical field, facilitating their directional movement. Macro-perspectives were correlated with direct and indirect oxidation processes [126]. During indirect oxidation, slag particles collide with cathodes and anodes, indirectly oxidizing vanadium and chromium as Fe³⁺ and H₂O₂ are generated.

3.2. Roasting-Leaching Processes

The roasting process of ash aims at the disintegration of the more stable and less soluble spinel phases, like Fe₂VO₄, as shown in Figure 8, and transforms them into more oxidized phases, enhancing the fact that they are more easily attacked by water/acid/alkaline solutions during the subsequent recovering process [128]. The roasting process is conducted with or without additives at temperatures ranging from 400 to 1000 °C, and advanced technologies have been utilized on fly ash to improve vanadium recovery [129,130]. To recover vanadium, the process involves oxidizing V³⁺ into acid-soluble V⁴⁺ and/or water-soluble V⁵⁺ compounds, which can then be dissolved using an appropriate reagent.

3.2.1. Non-Salt Roasting Assisted Leaching

This method results in the breakdown of the iron phases encasing the vanadium and chromium spinels, as explained in Equations (1) and (2) [62,108,132], allowing the formation of highly acidic leachable phases, including CaV₂O₆, Ca₃V₂O₈, MgV₂O₆, and Mg₃V₂O₈. As a result, the procedure involves the treatment of slags that have high levels of CaO/MgO.

4/5FeV₂O₄ + O₂ → 2/5Fe₂O₃ + 4/5V₂O₅ ΔG(T) = −293.45 + 0.15 T (kJ/mol O₂)

(1)

4/5FeCr₂O₄ + O₂ → 2/5Fe₂O₃ + 4/5Cr₂O₅ ΔG(T) = −47.01 + 0.04 T (kJ/mol O₂)

(2)

The olivine phase breaks down at 500 °C, whereas the spinels decompose at 800 °C, based on the results obtained earlier [36,110]. Wang [110] developed a method for direct extraction of low valence vanadium (LVV) from vanadium slag without producing poisonous waste, as shown in Figure 9. Application of roasting and leaching technique leads to the extraction of LVV from vanadium slag with a recovery rate of 69.37% at 800 °C roasting temperature for 1 h in an atmosphere of N₂/O₂ equals 10. Above this temperature or with longer roasting durations, V⁵⁺ was leached in the leaching solution. Smaller vanadium slag particles resulted in a higher vanadium recovery rate but lowered the amount of LVV.

In the same context, Vitolo [56] created their three-step technique by first burning the ash under controlled conditions to lower the amount of carbon in the ash (Figure 10). The recovery of vanadium increased to 97% in the subsequent leaching process because of the reducing the carbon component in the fly ash. The best-recommended temperature for the burning carbon phase was found to be 850 °C and resulted in a higher total recovery of V (83%) as V₂O₅ with fewer impurities. Lower V recovery resulted from higher burning temperatures (over 950 °C), which caused V to fuse and volatilize as well as produce complex V-Ni refractory compounds.

According to Li [109,133,134], the oxidation process of vanadium spinel can be described as follows: (i) conversion of vanadium spinel into a solid solution of Fe₂O₃.V₂O₃ at the early stages of roasting, (ii) oxidation of this solid solution to Fe₂O₃.V₂O₄, followed by an incomplete reaction of V⁴⁺ with MgO to produce Mg₂VO₄, and (iii) extended roasting time leads to the further oxidation of vanadium, resulting in the formation of the highest-valence vanadates. In the non-salt roasting process, the chromium spinel could not convert to carcinogenic chromate salts to avoid the cost and disposal of chromium waste. Also, the vanadium may be directly leached out as NH₄VO₃, which can be separated after cooling and crystallization (Figure 11). The leaching efficiency of vanadium could be up to 90%.

To perform the thermodynamic analysis of this process, we attempted to create a diagram, illustrating the relationship between standard Gibbs free energy change and temperature (∆G⁰-T), and the data required for each reaction were obtained from HSC Chemistry 9.3.0/Fact-Sage software [127]. The ∆G⁰ values for all reactions are negative throughout the temperature range of 400 to 1000 °C.

This indicates that direct oxidation of the spinels (FeV₂O₄/FeCr₂O₄) is thermodynamically feasible. The Gibbs free energy change for reaction (1.1) is much more negative than that of reaction (1.2), exhibiting that V-spinels can be preferentially oxidized and hence decomposed before that of Cr-spinels. Zhang [135] compared the different effects of microwave and conventional blank roasting on high-chromium vanadium slag’s oxidation conduct, microstructure, and surface structure. The normal spinel was oxidized to inverse spinel at 400 °C which then decomposed at 600 °C. Carrying out the roasting process at high temperatures resulted in the minority of Cr³⁺ ions in the spinel phase being incorporated into VO₂ to form the Cr_0.07V_1.93O₄ or CrVO₄. Wang [136] found that the roasting process depends on the original state of vanadium in the coal stone. In the case of existing vanadium in amorphous phase form, the non-salt-roasting technology is excellent for leaching out vanadium. While the occurrence of vanadium in the vanadium-bearing waste is in a crystalline phase, the addition of a fluxing agent is necessary to achieve high vanadium leaching efficiency.

Table 6. Various studies on vanadium’s non-salt roasting and leaching from various V sources.

Ref.	Non-Salt Roasting	Leaching		Results
	Conditions	Lixiviant	Conditions
[56]	850 °C, −250 µm for 60 min.	H2SO4	2.0 M of H2SO4 Conc., at 100 °C; for 60 min.	V = 97%
[36]	850 °C for 60 min.	Na2CO3	160.0 (g/L) of Na2CO3 Conc., and 10 (mL/g) L/S ratio at 95 °C for 150 min	V = 90% Cr remains in the residue
[109]	900 °C for 150 min.	(NH4)2CO3	25% of Na2CO3 Conc., and 4 L/S ratio at 50 °C for 150 min.	V = 92.8%
[110]	10 N2/O2, at 800 °C for 60 min.	H2SO4	2 M of H2SO4 Conc., at 90 °C; for 150 min.	V = 69.37%
[133]	900 °C for 150 min.	(NH4)2C2O4	13% of (NH4)2C2O4 Conc., and 4 L/S ratio at 70 °C for 60 min.	V = 90%

presents a summary of the extensive research conducted on vanadium resources through non-salt roasting and leaching procedures.

However, non-salt technology offers several advantages, such as reducing the carbon content of HOA, lowering its volume, and increasing the levels of V and Ni in the ash. There are some limitations, including the release of gas emissions, treating only waste with a high content of CaO or MgO, the chance of V compounds becoming volatile, the possibility of fusion and the creation of V-Ni refractory compounds at temperatures exceeding 900 °C, and alterations in the ash pH that might negatively impact the recovery of V and Ni during the leaching process.

3.2.2. Calcification Roasting Assisted Leaching

As an alternative to non-salt roasting, the clean technique of calcification roasting was used. The calcium roasting process involves combining limestone, lime, or other calcium compounds with vanadium wastes, grinding the mixture to an acceptable size, and then roasting the mixture in a vertical kiln [137,138], as illustrated in Figure 12. According to the chemical reactions listed in

Table 7. Reactions during calcium-roasting and leaching of vanadium-bearing waste.

Calcium Oxide (CaO)	∆G (kJ/mol O2)	Eqs.
4/5FeV2O4 + 4/5CaO + O2→4/5CaV2O6 + 2/5Fe2O3	∆G(T) = −439.86 + 0.21T (kJ/mol O2)	(3)
4/5FeV2O4 + 12/5CaO + O2→4/5Ca3V2O8 + 2/5Fe2O3	∆G(T) = −655.92 + 0.44T (kJ/mol O2)	(4)
4/7FeCr2O4 + 8/7CaO + O2→8/7CaCrO4 + 2/7Fe2O3	∆G(T) = −255.77 + 0.15T (kJ/mol O2)	(5)
2/3MgCr2O4 + 4/3CaO + O2→4/3CaCrO4 + 2/3MgO	∆G(T) = −468.37 + 0.32T (kJ/mol O2)	(6)
Acid-leaching of calcified roasted product
CaV2O6 + 2H2SO4→(VO2)2SO4 + CaSO4 + 2H2O		(7)
Ca3V2O8 + 4H2SO4→(VO2)2SO4 + 3CaSO4 + 4H2O		(8)
CaCrO4 + H2SO4→H2CrO4 + CaSO4		(9)
Alkali-leaching of calcified roasted product
CaV2O6 + Na2CO3→ 2NaVO3 + CaCO3		(10)
Ca3V2O8 + 3Na2CO3→2Na3VO4 + 3CaCO3		(11)
Ca3V2O8 + 3(NH4)2CO3→2(NH4)3VO4 + 3CaCO3		(12)
CaCrO4 + (NH4)2CO3→(NH4)2CrO4 + CaCO3		(13)

and the previous studies of thermodynamic assessment, the oxidation processes involving CaO and V-spinels (Equations (3) and (4)) begin around 600 °C, whereas those using Cr-spinels (Equations (5) and (6)) start at 800 °C [139,140]. It is noted that the vanadium and calcium reaction during the roasting process leads to the formation of calcium vanadates (CaV₂O₆ and Ca₃V₂O₈); the compounds mentioned are not soluble in water (Equations (7)–(13)); therefore, they must be leached out through H₂SO₄ acid (Figure 13) [137,141,142] or alkaline leaching with ammonium carbonate solution (Figure 14) [139,143,144].

Roasting and leaching behaviors of vanadium and chromium using calcification roasting-acid leaching from high-chromium vanadium slag were investigated [145]. The results indicated that more CaO combined with vanadium and reacted to form calcium vanadate (Ca₃V₂O₈), resulting in a high leaching efficiency of vanadium, approximately 91.14%, while the chromium remained in the leaching residue, and only 8.48% was leached. Gao [130] studied the roasting process of high chromium vanadium slag using CaO and microwave heating techniques. The effect of CaO dosage on vanadium-roasted products was investigated. At low m(CaO)/m(V₂O₅), CaV₂O₆ was formed, which was then converted to Ca₂V₂O₇ and CaVO₃ with increasing m(CaO)/m(V₂O₅) ratio. Results indicated that microwave radiation can reduce particle size and shorten roasting time, achieving 98.29% vanadium leaching efficiency under optimal conditions. Zhang [42] investigated the use of lime in roasting vanadium-bearing waste. Vanadium recovery was significantly influenced by the heating rate. Hence, reducing the heating rate resulted in high vanadium recovery. The formation of Ca₂V₂O₇ was more favorable for leaching vanadium with sulfuric acid. Xiang [146] applied mechanical activation treatment to enhance the extraction of vanadium from converter slag at a 1:1 mole ratio of vanadium to calcium. Then, the roasted slag was leached in a 15% sulfuric acid solution for 60 min at a temperature of 50 °C, stirring speed of 150 rpm, and S/L ratio of 1:20. The results demonstrated that the mechanical activation significantly decreased the optimum roasting temperature from 900 to 800 °C and increased the leaching efficiency of vanadium from 86.0 to 90.9%, respectively.

Table 8. Results of different studies on vanadium’s calcium roasting and leaching from various V sources.

Ref.	Roasting		Leaching		Results
	Salt	Conditions	Lixiviant	Conditions
[143]	CaO	1:1.1 V/Ca molar ratio, 900 °C for 120 min.	(NH4)2CO3	600 g/L of (NH4)2CO3 Conc., 20 (mL/g) L/S, 80 °C, 70 min.	V = 96% P = 9.2%
[144]	CaO	900 °C for 180 min.	NH4HCO3	15% NH4HCO3, 75 °C, 180 min.	V = 69.2%
[147]	CaO	1:16 CaO/solid ratio, 850 °C for 120 min.	H2SO4	15% H2SO4 Conc., 10 (mL/g) L/S ratio, 55 °C, 70 min.	V = 93%
[42]	CaO	0.42 CaO/solid ratio, 850 °C for 150 min.	H2SO4	pH 2.5, 4 L/S ratio, 65 °C, 60 min.	V = 91.5%
[148]	CaO	0.5 CaO/V2O3 molar ratio, 900 °C for 60 min.	H2SO4	20% H2SO4 Conc., 5 (mL/g) L/S ratio, 50 °C, 60 min.	V = 99.4%
[149]	CaO	1 CaO/V2O5 molar ratio, 900 °C for 120 min.	(NH4)2SO4 + H2SO4	250.0 g/L (NH4)2SO4 Conc., 3.75 M H2SO4 conc., 10 (mL/g) L/S ratio, 20 °C, 60 min.	V = 93.5% Cr = 0.2%
[139]	CaO	0.5 CaO/V2O5 molar ratio, 900 °C for 60 min.	Na2CO3	160 (g/L) Na2CO3 Conc., 10 L/S, 80 °C, 60 min.	V = 93.2% Cr = 0.04%
[150]	CaCO3	1 Ca/V molar ratio, 850 °C for 120 min.	H2SO4	15% H2SO4 Conc., 10 (mL/g) L/S, 10 (mL/g), 50 °C, 60 min.	V = 83%
[151]	(1st) CaO (2nd) Na2CO3	1st stage: 5 CaO/V2O3 molar ratio, 780 °C for 60 min 2nd stage: 3.3 Na2CO3/Cr2O3 at 950 °C	1st stage: H2SO4 2nd stage: Water	1st stage: 3.75 M H2SO4 Conc., 5 L/S ratio, 70 °C, 60 min. 2nd stage: 3 L/S ratio, 3 (mL/g), 25 °C, 20 min.	1st stage: V = 90% Cr = 1% 2nd stage: Cr = 87% V = 99%

summarizes extensive research on different vanadium resources utilizing calcium roasting and leaching techniques. Utilizing this technology makes it possible to obtain high-purity vanadium leach solutions with minimal interference from elements such as silicon and phosphorus. The other advantage is that the leaching may proceed at a lower temperature when a particular ammonia salt is used. However, the leaching efficiency is much lower when using ammonium salts compared to the maximum 90% filtration of vanadium with sodium salts due to the precipitation of vanadium as NH₄VO₃. So, problems linked to ammonia volatilization must also be considered. Additionally, there are several drawbacks associated with this technology, such as the production of a large quantity of leaching residues containing sulfate compounds, equipment corrosion, and the emission of harmful gasses, which can cause environmental contamination.

3.2.3. Sodium Salt Roasting Assisted Leaching Process

Traditional roasting technologies include sodium roasting, which can be traced back to Bleecker’s first salt-roasting technology in 1912 [128]. Basically, a mixture of sodium salts (Na₂CO₃, Na₂SO₄, NaCl, and NaOH) as a source for alkalis was roasted with vanadium-bearing waste in a furnace at high temperatures, depending on the melting points of the sodium sources as follows: 1200~1250 °C for Na₂SO₄, 750~850 °C for NaCl, 800~1000 °C for Na₂CO₃, and 400~800 °C for NaOH [43,152,153]. A maximum amount of oxygen was used during roasting to oxidize the V³⁺ and convert it into soluble sodium vanadates (NaVO₃, Na₄V₂O₇) through the chemical reactions listed in

Table 9. Reactions during salt roasting and leaching of vanadium-bearing waste.

Sodium Sulfate (Na2SO4)	∆G (kJ/mol O2)	Eqs.
4/3FeV2O4 + 4/3Na2SO4 + O2→8/3NaVO3 + 2/3Fe2O3 + 4/3SO2	∆G(T) = −155.62 − 0.08T	(14)
4FeV2O4 + 8Na2SO4+ O2→4Na4V2O7 + 2Fe2O3 + 8SO2	∆G(T) = 985.43 − 1.19T	(15)
FeCr2O4 + 2Na2SO4 + O2→2Na2CrO4 + 1/2Fe2O3 + 2SO2	∆G(T) = 936.43 − 0.59T	(16)
2MgCr2O4 + 4Na2SO4 + O2→4Na2CrO4 + 2MgO +4SO2	∆G(T) = 1195.64 − 0.78T	(17)
Sodium chloride (NaCl)
4/5FeV2O4 + 8/5NaCl + 4/5H2O + O2→8/5NaVO3 + 2/5Fe2O3 + 8/5HCl	∆G(T) = −245.36 + 0.04T	(18)
4/9FeV2O4 + 16/9NaCl + 4/9H2O + O2→4/9Na4V2O7 + 2/9Fe2O3 + 8/9HCl	∆G(T) = −113.65 − 0.05T	(19)
4/7FeCr2O4 +16/7NaCl + 8/7H2O + O2→8/7Na2CrO4 + 8/7Fe2O3 + 16/7HCl	∆G(T) = 79.07 − 0.01T	(20)
2/3MgCr2O4 + 8/3NaCl + 4/3H2O + O2→4/3Na2CrO4 + 4/3MgO + 8/3HCl	∆G(T) = 148.14 − 0.12T	(21)
4/7FeV2O4 + 8/7NaCl + O2→8/7NaVO3 + 2/7Fe2O3 + 4/7Cl2	∆G(T) = −186.26 + 0.09T	(22)
4/9FeV2O4 + 16/9NaCl + O2→4/9Na4V2O7 + 2/9Fe2O3 + 8/9Cl2	∆G(T) = −80.26 + 0.06T	(23)
4/7FeCr2O4 + 16/7NaCl + O2→8/7Na2CrO4 + 2/7Fe2O3 + 8/7Cl2	∆G(T) = 55.83 + 0.02T	(24)
2/5MgCr2O4 + 8/5NaCl + O2→4/5Na2CrO4 + 2/5MgO + 4/5Cl2	∆G(T) = 36.74 + 0.005T	(25)
Sodium carbonate (Na2CO3)
4/5FeV2O4 + 4/5Na2CO3 + O2→8/5NaVO3 + 2/5Fe2O3 + 4/5CO2	∆G(T) = −345.3 + 0.04T	(26)
4/5FeV2O4 + 8/5Na2CO3 + O2→4/5Na4V2O7 + 2/5Fe2O3 + 8/5CO2	∆G(T) = −306.70 + 0.07T	(27)
4/7FeCr2O4 + 8/7Na2CO3 + O2→8/7Na2CrO4 + 2/7Fe2O3 + 8/7CO2	∆G(T) = −94.64 − 0.04T	(28)
2/3MgCr2O4 + 4/3Na2CO3 + O2→4/3Na2CrO4 + 2/3MgO + 4/3CO2	∆G(T) = −92.39 + 0.14T	(29)
Sodium hydroxide (NaOH)
4/5FeV2O4 + 8/5NaOH + O2→8/5NaVO3+ 2/5Fe2O3 + 4/5H2O	∆G(T) = −458.66 + 0.11T	(30)
4/5FeV2O4 + 16/5NaOH + O2→4/5Na4V2O7 + 2/5Fe2O3 + 8/5H2O	∆G(T) = −533.41 + 0.08T	(31)
4FeCr2O4 + 8NaOH→4Na2CrO4 + 2Fe2O3 +4H2O	∆G(T) = −273.10 − 0.02T	(32)
2/3MgCr2O4 + 8/3NaOH + O2→4/3Na2CrO4 + 2/3MgO + 4/3H2O	∆G(T) = −210.17 + 0.01T	(33)
Water leaching of sodium roasted product
${\mathrm{NaVO}}_3+{\mathrm{H}}_2\mathrm{O}\to {\mathrm{H}}_2\mathrm{VO}{\displaystyle \genfrac{}{}{0pt}{}{-}{4}}$ + Na+		(34)
${\mathrm{Na}}_4{\mathrm{V}}_2{\mathrm{O}}_7+{\mathrm{H}}_2\mathrm{O}\to 2\mathrm{HVO}{\displaystyle \genfrac{}{}{0pt}{}{2-}{4}}$ + Na+		(35)
Acid-leaching of sodium roasted product
2NaVO3 + 2H2SO4→(VO2)2SO4 + Na2SO4 + 2H2O		(36)
Na4V2O7 + 3H2SO4→(VO2)2SO4 + 2Na2SO4 + 3H2O		(37)
Na2CrO4 + H2SO4→H2CrO4 + Na2SO4		(38)
Alkaline-leaching of sodium roasted product
Na2O.V2O5 + 4NaOH→2Na3VO4 + 2H2O		(39)
2NaVO3 + (NH4)2CO3→2NH4VO3 + Na2CO3		(40)
Na4V2O7 + 2(NH4)2CO3→(NH4)4V2O7 +2Na2CO3		(41)
Na2CrO4 + 2(NH4)2CO3→(NH4)4V2O7 + 2Na2CO3		(42)

[127]. The vanadium conversion degree in these sodium sources follows the following order: NaOH > Na₂SO₄ > Na₂CO₃ > NaCl [154]. This order can be attributed to several reasons, including the diffusivity of the different sodium sources into the interior layers of the wastes, where they react with the V-spinels [155,156,157]. Despite all the limitations of this technology, which include the formation of harmful gaseous products, such as SO₂, Cl₂, or CO_2, fusion of NaCl at high temperatures, and high energy consumption while using sodium salts as the alkali source, it remains the best method in terms of selective vanadium extraction with low operating cost. The above serious environmental pollution greatly hinders the sustainable development of V extraction from V-bearing materials [158,159]. Consequently, efforts are still needed to develop an eco-friendly and effective method for extracting V from V-bearing materials. Attention has also been paid to recycling the Na⁺ from the leach solution, using a membrane-assisted electrochemical cell to separate the cation from vanadium ions [160].

Anyway, the water-leaching process mostly applies to products undergoing sodium roasting, which takes place through Equations (34) and (35). Metal sulfates are formed by dissolving sodium vanadate compounds by H₂SO₄, according to Equations (36)–(38), and ammonium carbonate leaching is accomplished through Equations (39)–(42). However, the leaching effectiveness depends on the degree of metal converted during roasting, where sodium vanadate can be leached out. With regard to this case, there are certain common features for the processes that are used in the commercial plants of vanadium production from industrial wastes, such as the sodium salt roasting-water leaching-ammonia precipitation process utilized at Highveld Steel and Vanadium Corp. (South Africa), Chengde Iron and Steel Group (China), Nizhniy Tagil Iron and Steel Works (NTMK, Russia), etc. [127]. Li [43] reported that only 87.9% of V and 6.3% of Cr were recovered from calcined waste at Na/V molar ratio of 3.3 at 800 °C using water leaching, as shown in Figure 15. When the leaching residue was subjected to second sodium calcination at (Na/(V, Cr) molar ratio of 2.86 and 950 °C), about 90.7% of vanadium and 96.4% of chromium were leached, giving an overall recovery of vanadium and chromium of 98.9% and 96.4%, respectively.

Wen [161] evaluated the leaching of the vanadium and chromium using sodium salts roasting (Na₂CO₃) and then leaching them with (NH₄)₂SO₄, showing that 94.6% vanadium and 96.5% chromium were leached. According to Zhao [162], vanadium can be leached from stone coal with the addition of 6 wt.% sodium chloride and 10 wt.% sodium sulfates during the roasting process. It was noted that the vanadium-bearing muscovite with quartz was converted to feldspar group minerals (albite, orthoclase, and anorthite). These results could produce some potential methods for vanadium recovery from stone coal. Puhong [129] roasted red cake obtained from multi-processing of stone coal with 22.5 g NaOH/25 g red cake at 170 °C for 1 h, then used water leaching at 98 °C for 60 min. with a solid/liquid ratio of 1:3.3 g/mL. It was noted that the V leaching efficiency was up to 97% after purification and calcination as V₂O₅, with a purity of 99.3%.

A three-step process was performed on HOA containing 85 wt.% unburned C and 2.2 wt.% V, beginning with carbon burning, followed by salt roasting and water leaching. It was noted that the 4 h. calcining step at 650 °C removes most of the carbon and thus increases the V content to 19 wt.%. Then, the burned enriched-V ash was roasted with sodium carbonate Na₂CO₃ at 650 °C for 4 h to convert V oxides into sodium metavanadate (NaVO₃), which is a water-soluble compound [163]. The calcined ash was leached with water for 4 h at 60 °C with an S/L ratio of 1:50 g/mL. Water leaching leads to the selective dissolving of V with about 92% recovery, leaving Fe and Ni compounds undissolved in the residue of the ash. By adding (NH₄)₂SO₄, V is precipitated as NH₄VO₃ from the leaching solution to separate V from the solution [101].

In the same context, Ibrahim et al. [59] discussed a recent technique for converting V compounds into water-soluble vanadates (NaVO₃) through roasting with NaCl, as shown in Figure 16. They found that over 95.5% of the vanadium leached out after the HOA was roasted with 20% NaCl for 2.5 h. Then, the vanadium was leached using distilled water with an L/S ratio of 10 mL.g⁻¹. The purity of the obtained NH₄VO₃ powder was estimated to be about 92% using direct precipitation from vanadium-pregnant solution by NH₄Cl addition, and the precipitation efficiency of NH₄VO₃ powder was estimated to be around 91.5%, and the total recovery rate of vanadium reached up to 87.60%. Then, the ammonium meta-vanadate was calcined at 550 °C for 3 h to produce V₂O₅ powder with an estimated 83% purity (Figure 17).

The authors [59] comprehensively investigated the solid-state and phase transformation mechanisms that govern the selective extraction of vanadium from HOA. The SEM images are displayed in Figure 18 for the roasted sample produced at 700 °C, and the corresponding EDS analysis for different points on the burnt ash is listed in

Table 10. Spots EDS analysis of roasted boiler ash at 700 °C for 2 h and 20 wt.% of NaCl.

Spots	Composition (wt., %)
	V	Ni	Fe	O	Na	Mg	Al	Si
S1	18.57	2.93	5.88	28.07	25.39	15.12	3.03	1.01
S2	2.24	7.72	4.47	26.34	23.21	31.59	3.15	1.28
S3	1.2	10.39	17.58	20.56	18.05	22.88	5.61	3.73
S4	10.08	4.61	6.01	30.5	28.24	13.5	6.02	1.04
S5	3.94	3.14	25.2	24.6	19.83	11.42	9.34	2.53

. The results (Figure 18a) showed that the roasted particles are presented in an irregular shape with different sizes, and a few areas of spinel surface structure have been destroyed. According to the elemental mapping analyses shown in points S₁ and S₄, the decomposition of the spinel started to release the vanadium molecules, which indicates that sodium and vanadium are partially compatible (Figure 18b). The main information shown in this figure includes the following: (1) the particles are loose and dispersed, indicating that the interaction between particles is not significant under this temperature; (2) the distribution of sodium is relatively extensive and uniform, indicating that NaCl has not entirely reacted and formed a new sodium phase; and (3) the relative enrichment of vanadium appeared in some particles with high sodium content, indicating that some phases containing sodium and vanadium were formed.

Based on the EDS analysis (Table 10) of the specific locations (S₁ and S₄), it was clear that the chemical composition showed a moderate correlation between the levels of vanadium and sodium due to the limited reaction range of O₂ with vanadium spinel to release their oxides. So, not all quadruple vanadium (V⁴⁺) was oxidized to the highest V⁵⁺ valency [64,164]. As a result, the reaction temperature at 700 °C was not sufficient to convert more of the V found in the vanadium compound into water-soluble sodium vanadates because almost all of V still exists inside the spinel, with less outer distribution than the central region and not enough opportunity to react with NaCl, resulting in decreasing the vanadium leachability [111]. To further explore, a line scan analysis was performed, as shown in Figure 19. Clearly, it was noted that the V and Na lines are aligned to some extent, suggesting an inconsiderable liberation of the vanadium element from the spinel surface. Furthermore, a large amount of free sodium remains along the line, meaning it did not enter into the chemical reaction.

Their findings indicate that a roasting temperature of 850 °C (Figure 20) appears optimal for facilitating the O₂ reaction with vanadium-bearing minerals. This temperature provides an excellent opportunity for vanadium molecules to react with sodium chloride, forming the sodium metavanadate phase (NaVO₃). However, this phase can be easily extracted through a water-leaching process, resulting in the liberation of vanadium molecules, as shown in. Elemental scan lines demonstrated that the Na and V lines align, indicating that a considerable amount of the V element has been released from the spinel surface.

Elemental scan lines demonstrated the distribution of elements in the roasted ash at 850 °C is shown in Figure 21. Also, it was noted that there is an abroad alignment of the Na and V lines, suggesting that there is a significant liberation of the V element from the spinel surface. Also, along the line, there are no remains of free sodium, which means that all of it was entirely consumed in the process of forming a water-soluble sodium vanadate compound. This result confirms the hypothesis explained in the TGA data of the same study, which has been attributable to the complete thermal decomposition of vanadium spinel (FeV₂O₄) and vanadyl sulfates at 850 °C.

Furthermore, there are no signs of unbound sodium present in the mixture, suggesting that it was entirely consumed in the process of forming a water-soluble sodium vanadate compound. The results of the point-scan analysis of three different surface products are listed in

Table 11. Spots EDS analysis of roasted boiler ash at 850 °C and 1000 °C for 2 h and 20 wt.% of NaCl.

	Spots	Composition (wt., %)
		V	Ni	Fe	O	Na	Mg	Ca	Al	Si
850 °C	S1	17.86	1.45	5.29	38.23	20.45	7.19	1.02	3.96	4.55
	S2	12.62	1.74	4.83	35.01	25.11	7.05	0.98	7.58	5.08
	S3	10.06	1.41	3.27	37.54	27.04	6.89	1.08	7.26	5.45
	S4	14.08	1.97	2.46	33.93	25.06	8.87	1.5	5.97	6.16
	M1	0.92	22.11	38.77	27.87	1.15	5.45	1.01	1.63	1.09
	M2	0.81	4.28	7.24	35.3	15.42	3.95	1.92	14.57	16.51
	M3	1.07	19.95	36.81	25.76	2.08	9.03	1.02	1.45	2.83
1000 °C	S1	19.35	16.76	23.21	25.61	1.03	9.5	2.02	1.51	1.01
	S2	18.87	15.98	21.28	27.84	1.31	10.46	1.98	1.36	0.92
	S3	19.06	19.19	20.14	26.75	1.56	9.69	1.04	1.47	1.1
	M	-	-	-	47.18	3.14	32.21	1.5	-	15.97
	N1	-	-	-	30.65	46.04	-	3.51	9.7	10.1
	N2	-	-	-	28.71	39.45	-	2.08	19.84	9.92

. Meanwhile, the excessive roasting temperature of up to 1000 °C (Figure 22) would allow for the formation of refractory compounds (i.e., sintered from V, Ni, Fe). In addition, the formation of the aluminum silicate (NaAlSi₂O₆) phase increases due to the melting of NaCl salt, which leads to the formation of agglomerate that could hinder oxygen transfer during the roasting process and hence decrease the vanadium percent recovery.

Generally, the direct leaching process is poorly selective and has varying leaching efficiencies among different vanadium-containing phases. Consequently, subsequent separation and purification processes for vanadium and other metals in the leaching solution are more complex, with a wide range of reagent types and high consumption rates. On the other hand, the roasting-leaching process enables control over the occurrence form of vanadium through roasting, thereby ensuring efficient vanadium leaching. Specifically, the sodium salt roasting process can be employed to selectively separate vanadium using the water leaching method, which offers advantages such as high leaching efficiency, excellent selectivity, reduced consumption of leaching reagents, and suitability for treating materials with complex forms of vanadium occurrence. Based on the characteristics of V-bearing waste of high vanadium content and complex vanadium occurrence forms, NaCl has been used, a low-cost and abundantly available salt compared to other salts in the roasting-water leaching process, to study the extraction process of vanadium and focus on the transition law and mechanism of the vanadium-containing phase during sodium salt roasting so as to provide the theoretical basis for efficient extraction of vanadium. A summary of the comprehensive research carried out on various vanadium resources using sodium-salt roasting and leaching processes is presented in

Table 12. Results of different studies on vanadium’s sodium-salt roasting and leaching from various V sources.

Ref.	Roasting		Leaching		Results
	Sodium-Salt Roasting	Conditions	Lixiviant	Conditions
[59,165]	NaCl	1st stage Roasting at 850 °C for 150 min. 2nd stage Water leaching solid residue (WLSR)	Water H2SO4	10 (mL/g) L/S ratio at 25 °C for 90 min. −63 µm, 8% H2SO4 Conc., in 15 (mL/g) L/S ratio at 85 °C; for 240 min.	1st stage: V = 95.5% 2nd stage: Ni = 95.02% Zn = 90.13%
[43]	Na2CO3 + O2	1st stage (V-roasting) 3.3 Na/V molar ratio at 800 °C for 120 min. 2nd stage (Cr-roasting) 2.86 Na/(V,Cr) molar ratio at 950 °C for 120 min.	Water	1st stage: 3 (mL/g) L/S ratio at 25 °C; for 20 min. 2nd stage: 3 (mL/g) L/S ratio at 25 °C; for 20 min.	1st stage: V = 87.9% Cr = 6.3% 2nd stage: Cr = 96.4% V = 90.7%
[166]	Na2CO3	41/9 Na2CO3/solid ratio at 700 °C for 150 min.	Water	5 (mL/g) L/S ratio at 90 °C for 30 min.	V = 90%
[167]	Na2CO3	0.1 Na2CO3/solid ratio at 0.2 1000 °C for 45 min.	Na2CO3 + NaOH	45 + 10 (g/L) of Na2CO3 + NaOH Conc., in 20 L/S at 80 °C for 60 min.	V = 80%
[41]	Na2CO3	1.0 Na2CO3/solid ratio, at 2.0 1000 °C for 120 min.	H2SO4	3.0 M H2SO4 Conc., in 15 (mL/g) L/S ratio at 70 °C; for 150 min.	V = 95%
[161]	Na2CO3	2.5 Na2CO3/solid ratio at 850 °C for 60 min.	(NH4)2SO4	30.0 (g/L) (NH4)2SO4 Conc., in 3 (mL/g) L/S ratio at 20 °C for 60 min.	V = 94.6% Cr = 96.5%
[153]	NaOH + O2	7.67 Na/V molar ratio at 700 °C for 120 min.	Water	70 (mL/g) L/S ratio at 25 °C for 60 min.	V = 99%
[168]	NaOH	0.3 NaOH/solid ratio at 800 °C for 180 min.	Water	10 (mL/g) L/S ratio at 30 °C for 60 min.	V = 94.9% Cr = 80.5%
[169]	Na2CO3 + NaCl	3:2:5 Na2CO3/NaCl/solid ratio at 700 °C for 120 min.	Water	4 (mL/g) L/S ratio at 95 °C for 180 min.	V = 96% Cr = 91%
[170]	NaOH	0.3 NaOH/solid ratio at 850 °C for 120 min.	NaOH pressure leaching	5 (mL/g) L/S ratio; 0–5 MPa O2 partial pressure at 210 °C for 120 min.	V = 98%

.

Likewise, Ibrahim [165] studied the cost-effective extraction of precious metals, such as Ni and Zn, by hydro-metallurgical processing of the water-leaching solid residue (as shown in Figure 23) after vanadium extraction from salt-roasting Egyptian boiler ash [59]. Under the most favorable leaching conditions of 8% (vol%) H₂SO₄ concentration, 85 °C leaching temperature, and 1/15 S/L ratio, a maximum extraction of 95.02% Ni and 90.13% Zn was achieved after 240 min of leaching, while the iron impurity was removed by Fe₂O₃ as a nucleating agent and the magnesium impurity was effectively removed by oxalic acid precipitant. After the removal of Fe and Mg, Ni and Zn in the purified solution were precipitated at a pH of 10 as Ni-Zn hydroxide (Ni(OH)₂ and Zn(OH)₂), which was subsequently transformed into NiO-ZnO by its calcining at 450 °C for 2 h. The precipitation efficiency of Ni and Zn was 95.25% and 89.51%, respectively, and the final calcined product was composed mainly of 37% Ni and 23% Zn. Based on a kinetic analysis, it was discovered that the process of leaching nickel is primarily controlled by diffusion through the solid product layer and chemical reactions. The diffusion process through the solid product layer is the main contributor, with an activation energy of 20.26 kJ/mol. The kinetic of zinc dissolving is governed by the diffusion that occurs through a layer of a solid product, and this diffusion has an activation energy of 11.67 kJ/mol.

3.2.4. Promising Modification Methods

Continuous efforts have been made to develop alternative processes, including sub-molten salts and supercritical fluid for recovering vanadium from various sources. Sub-molten salt and novel leaching methods can extract vanadium from various sources (vanadium-bearing slag, stone coal, spent catalysts, etc.) and should be discussed in this review. The method of sub-molten salts (NaOH/KOH/NaOH-NaNO₃ = 75% (w/w) at high pressure) can decrease the cost of energy due to the lower reaction temperature [171,172]. According to a report by Liu [171], a new study has discovered a unique approach to breaking down vanadium slag using KOH as a sub-molten salt under regular pressure. When the reaction conditions are optimal, including a temperature of 180 °C, an initial ratio of 4:1 potassium hydroxide to ore mass, a stirring speed of 700 rpm, a gas flow rate of 1 L/min, and a reaction time of 300 min, the extraction rates of vanadium and chromium can attain up to 95% and 90%, respectively. Still, the requirement of reactors with high corrosion resistance, the large consumption of alkaline, and the high dissolution rate of silica phases limit its wide application. The problems of inefficient vanadium separation and serious environmental pollution greatly hinder the application of these V extraction processes [158,159]. Recently, novel leaching methods have been reported for vanadium recovery using urea, specific chelating agents, or supercritical fluid, resulting in low environmental impact [173,174,175,176].

A leaching method has also been reported by adjusting the vanadium particles’ surface wettability to intensify the solid–liquid contact using certain surfactants [177]. Though tested in a few cases, as shown in

Table 13. Details of some novel leaching methods for vanadium recovery from various sources.

Source	Method	Conditions	Results	Principle	Highlight	Ref.
Vanadium spent catalyst	Urea leaching	Time, 1 h; temperature, 20 °C; urea conc., 20%; pH = 14	V = 80%	Considering the mutual solubility effect of urea to vanadium solubility in water	High kinetic leaching of vanadium even at low temperatures is an advantage	[175]
Vanadium-bearing shale, vanadium-bearing coal	Chelating leaching	Time, 4 h; temperature, 95 °C; chelator conc., 6.0 M; additive (CaF2), 5% (wt.)	87% V	The principle is similar to that of bioleaching, which is to obtain a product of a complex metal-chelator. Chelators used are citric acid, oxalic acid, and ethylenediaminetetraacetic acid (EDTA).	The process offers the advantage of an eco-friendly approach compared to conventional leaching using mineral acids. Even so, slower kinetic leaching becomes the main problem of this approach, being the limitation for its wider application.	[3,173]
Vanadium-bearing slag	Surface wettability control	Leaching conditions: not specified Surfactant added: 0.25–1.0% (wt.)	69% V (with surfactant), 51% V (without surfactant)	To intensify the interaction between the solid particle and leaching solution by adding surfactants like polydimethylsiloxane/PDMS and sodium dodecyl sulfate/SDS. The surfactants play roles in reducing the surface tension of the solution and changing the surface properties of the solid particles	The pioneered study showed the capacity of this approach to enhance vanadium extraction. Further studies in this area are worth encouraging while assessing the economic and environmental impact of the added surfactants.	[177]
Vanadium spent catalyst	Supercritical leaching	Fluid, acetylacetone; temperature, 190 °C; time, 7 h	V = 60%	Characterized by the use of fluids where they are in their critical point; for instance, CO2 (critical pressure/Pc, 7.38 MPa; critical temperature/Tc, 31.1 °C), ethanol (Pc, 6.3 MPa; Tc, 241 °C), methanol (Pc, 7.85 MPa; Tc, 240 °C), acetylacetone (Pc, 4.05 MPa; Tc, 339 °C) and other alcohol groups.	Research on this area is still green. The advantages in the environment, energy, and chemical aspects are worth considering while preliminary studies show an unsatisfactory result. The development of novel supercritical solvents and co-solvents is believed to be the main factor attributing to the greater interest in applying this technology in the extractive metallurgy field, including extracting vanadium.	[178]

, such conditions have resulted in considerably lower vanadium recovery yields (60~70%); however, they can still be considered a “green option” to the traditional processes. Accordingly, the broadly available theoretical underpinnings and technological background of such approaches make them worth investigating in the future. Once proven promising, they may eventually be considered for exploitation and optimization.

4. Utilizing By-Products for Industrial Applications

The world is dealing with water scarcity issues due to the discharge of partially treated or untreated wastewater from industries and groundwater pollution. The demand for clean water has increased to meet the needs of a sustainable society. The need for cost-effective and efficient water treatment methods has become a growing concern as we look towards the future. This has led to an increased focus on recycling industrial waste to produce low-cost adsorbents and reduce environmental pollution [179]. Thus, developing an affordable, accessible, and highly effective adsorbent for removing hazardous metal ions presented in wastewater has become necessary. The utilization of solid industrial wastes by reusing them in the manufacturing of useful materials used for environmental purposes has been the subject of several recent studies, and it is one of the most crucial issues for maintaining sustainable development [180,181].

The solid wastes containing silica (SiO₂) and alumina (Al₂O₃), including industrial waste forms, i.e., fly ash, steel slag, waste perlite, waste porcelain, lithium slag, waste metallic residues, paper sludge, cupola slag, kaolin waste, windshield waste, asbestos wastes, etc., could be converted into ecological zeolite-based adsorbents [182,183,184,185,186,187,188,189,190,191]. To date, several physicochemical and solvothermal techniques, including the hydrothermal approach [192,193,194], alkali-fusion processes [195,196,197], sol–gel processes [198,199,200], microwave processes [201,202,203,204], and alkali-leaching method [205,206,207,208] have been adopted and developed to produce synthetic adsorbents. Figure 24 shows the major preparation processes and methods for synthesizing zeolite from industrial solid waste. So, in this review, we focused on the synthesis of low-cost adsorbent from industrial waste.

However, many researchers have studied the synthesis of zeolites from solid wastes to be used as an adsorbent for removing heavy metal ions from wastewater. Zeolites are crystal formations built on stiff anionic alumino-silicate structures with distinct pores or channels that link at cavities or cages [209]. These materials are favorable for adsorption processes due to their high cation exchange capacity (CEC), large surface area, good thermal stability, porosity, surface active functional groups, and nontoxicity [210,211,212]. New zeolite materials have been processed from fly ash, successfully eliminating heavy metal ions from aqueous solutions [213]. Chen [214] synthesized zeolite X by an alkaline fusion-hydrothermal reaction using industrial lithium slag containing SiO₂:Al₂O₃ in a weight ratio of 71.73:25.16 (wt. %). The results highlight the performance similarity between Zeolite NaX-1 and commercial Zeolite X.

Figure 24. Diagram of the different procedures for synthesizing zeolite from fly ash [215].

Figure 24. Diagram of the different procedures for synthesizing zeolite from fly ash [215].

R. Anuwattana [216] has reported successfully synthesizing Na-A type zeolite using industrial waste materials such as the solid by-product of cupola slag, which contains SiO₂/Al₂O₃ in a weight ratio of 47:12 (wt.%), and aluminum sludge from an aluminum plating plant. In their investigation, two preparation methods were carried out using the same starting material compositions. To begin the process, alkaline fusion was utilized, and then the material was subjected to hydrothermal treatment to produce sodium aluminosilicate. This substance was subsequently crystallized in a NaOH solution at a temperature of 90 °C for 1~9 h with varying H₂O/SiO₂ ratios. The results indicated that a more excellent ratio of H₂O/SiO₂ resulted in an increased crystallization rate. The highest level of crystallization was observed for Na-A after 3 h. The second method involved alkaline hydrothermal treatment without fusion, using the same conditions as the first procedure. However, this approach did not yield any Na-A zeolite. It is worth considering blast furnace slag (BFS) as a possible source of raw materials for zeolite synthesis. This material primarily comprises CaO, SiO₂, Al₂O₃, and MgO, with minor Fe, Ti, and Mn quantities. Several groups have reported that zeolite can be made from industrial waste.

However, prior attempts to use BFS as a chemical source of zeolite were unsuccessful due to its complex composition, specifically its high Ca content [192]. Hence, further study is needed to produce a variety of zeolites for possible uses. In the same context, Kuwahara [217,218] discovered a new way to synthesize a hydroxyapatite-zeolite micro composite from BFS in a recent study. The process involves using affordable chemical reagents such as H₃PO₄ and NaOH and suitable preparation techniques. This innovative approach provides practical solutions for waste management. Furthermore, using the alkali fusion process, Takaaki Wajima [197] demonstrated how to create zeolite-A, zeolite-X, and hydroxysodalite from the BFS. A more recent approach (Figure 25) for synthesizing a high-crystallin and affordable zeolite from a combination of SASR-kaolin as a readily accessible and low-cost raw material using an alkaline fusion hydrothermal process [219]. The weight ratio of the SASR- kaolin mixture of 1:1.5 gives the best composition and properties of the synthesized zeolite. The optimal conditions for the synthesized zeolite are a fusion temperature of 600 °C, a 1:3 wt. ratio (SASR-kaolin)-NaOH, a 1:4 solid–liquid ratio, an 80 °C crystallization temperature, and a 24 h crystallization time.

Figure 26 illustrates the mineralogical structure of the produced synthesized zeolite by mixing pretreated SASR and kaolin mixtures at different weight ratios in the presence of NaOH (Mixture-NaOH mass ratio of 1:1.3). As shown in Figure 26A, the Faujasite zeolite (Na₂ Al₂Si_3.8O_11.63·8H₂O) has been formed as the major mineral phase from SASR-Kaolin weight ratios of 1:0, 1:2, 1:1.5, 1:1, and 2:1.

However, the intensity of its peaks increased with increases in the mass of kaolin, and a high rate of crystallinity (85.21%) can be obtained when the ratio of SASR to kaolin is 1:1.5, due to the stability of the zeolite composition. A percentage of sodalite zeolite (Na₈(AlSiO₄)₆(OH)₂.4H₂O) groups have also existed at an SASR-Kaolin weight ratio of 0:1. Figure 26B shows that the crystallinity of the synthesized zeolite decreased from 85.21% to 69.76% at 1:2 mixture-NaOH weight ratio. However, when the system’s alkalinity is too high, a decline in crystalline is caused. This can be attributed to one of the following two reasons: either the created zeolite spontaneously transforms into hydroxyl sodalite, which has more excellent thermodynamic stability [220], or the zeolite will dissolve in the hot alkali solution because of its metastable state and exposure to disturbances [180,221].

Table 14. Synthetic zeolites from different industrial wastes and their advantages and limitations.

Raw Material	Chemical Composition %	Advantages	Limitations	Synthesis Route	Zeolite Type	Ref.
BFS	CaO (40.1), SiO2 (34.58), A12O3 (14.78), MgO (5.29)	Availability, low cost, convenient preparation steps.	Higher CaO content.	Hydrothermal	A	[217,218]
Natural obsidian	80.04% SiO2, 12.27% Al2O3, 0.16% TiO2, 0.84% FeO, 0.18% MgO, 1.10% CaO, 3.14% Na2O and 3.04% K2O	Higher silica content, availability.	Still, limited studies did not mention any specific application. Due to the wide range of behavior changes concerning synthetic conditions.	Hydrothermal	Organic template-free EMT-type, natrolite, Za-gmelinite	[222,223]
RHA (Residue is rich in amorphous silica)	80% silica, Al2O3, iron oxide, CaO, MgO, sodium and potassium oxides, and others	Low cost, ultrafine size, highly porous, and chemically reactive.	Pre-treatment of RHA and waste glasses increase the cost.	Hydrothermal	ZSM5, T, Na-Y	[224,225]
Waste glass materials	SiO2 (63%), Al2O3 (18%), B2O3 (10%), and alkaline earth oxides	Excellent mechanical and thermal properties, free of harmful elements.	Problem in handling, need a substrate or base material for commercial use.	Hydrothermal	A	[198,199,226]
SASR	SiO2 26.57%, Al2O3 5.81%, Fe2O3 18.93%, TiO2 0.13%, MgO 0.04%, K2O 0.13%, Na2O 0.18%, P2O5 3.61%, LOI 17.88%	Availability, low cost, ultrafine size, convenient preparation steps.	Pre-treatment (Physical and Chemical) needs to be a source of Al.	Hydrothermal, Alkali fusion	Faujasite	[219]
Paper sludge ash	SiO2 35.9%, Al2O3 22.8%, CaO 33.2%, Na2O 0.6%, MgO 4.5%, Fe2O3 0.9%, TiO2 2.2%	Amorphous and crystalline phases formed by incineration; low temperature required.	Low abundance of Si and significant Ca content.	Hydrothermal	Na-P1	[227,228]
Waste stone cake	SiO2 38.9%, Al2O3 12.2%, CaO 6.9%, Na2O 1.4%, K2O 1.8%, MgO 2.7%, Fe2O3 3.1%, CO2 32.6%, SO3 0.1%, P2O5 0.1%	Specific pore sizes and large surface areas, high silica content.	Most SiO2 and Al2O3 are in crystalline phases; it is difficult to dissolve into an alkali solution, contains impurities.	Hydrothermal, Alkali fusion	zeolite-A, P, X and ZSM-5	[195]
Clay materials (kaolin, smectite)	SiO2 46.5%, Al2O3 41.18%, Fe2O3 0.19%, TiO2 0.13%, MgO 0.04%, K2O 0.13%, Na2O 0.18%, ZrO2 0.01%, SO3 0.15%, P2O5 0.03%, LOI 16.25%	Availability, convenient source for producing low silica zeolites like Y, and use of kaolin waste for zeolite synthesis reduces the cost of reagents.	High energy consumption processes like grinding, calcination, and fusion are required. Raw materials mining destroys the natural landscape.	Hydrothermal, Alkali fusion	NaA, mordenite, Faujasite, and, NaP	[229]
Coal fly ash	SiO2 38.3%, Al2O3 34.8%, CaO 11.0%, Fe2O3 8.1%, Others 7.8%	The main constituents are silica and alumina, which offer the potential to convert it to zeolite, producing low-price zeolite with high purity; and no harmful effect.	Effect of the impeller type and agitation during the hydrothermal treatment stage of the process.	Hydrothermal, microwave-assisted hydrothermal method, and fusion methods.	X, Na-P1, A, Y	[183,206,230]
Lithium slag	SiO2 70.67%, Al2O3 27.24%, Fe2O3 0.52%, SO3 0.45%, CaO 0.29%, K2O 0.22%, MgO 0.16%, Na2O 0.13%, P2O5 0.12%, Others < 0.1%	High silicon aluminum ratio.	Very few experiments have been conducted in this field.	Hydrothermal	X, FAU/LTA	[214,231]
Waste of iron mine tailings/iron ore tailing	SiO2 67.58%, Al2O3 8.70%, Fe2O3 7.42%, CaO 5.78%, MgO 4.37%, K2O 2.32%, Na2O 2.15%, Cl 0.69%, TiO2 0.33%, P2O5 0.26%, SO3 0.23%, MnO 0.10%, SrO 0.06%	Have economic and environmental aspects.	Very few experiments have been conducted in this field, which contains hazardous impurities that decrease the quality of zeolites.	Hydrothermal	A, ZSM-5	[232]

illustrates how various industrial wastes can be used to create synthetic zeolites, each with its own advantages and disadvantages.

Additionally, in the same context of research [219], synthesized zeolite was an effective adsorbent for removing Zn²⁺, Pb²⁺, Cu²⁺, and Cd²⁺ heavy metal ions from industrial wastewater based on readily accessible and low-cost raw materials. It was noted that the adsorption process was exothermic, pH-dependent, and spontaneous in nature. Significant adsorption occurs when the pH is 7.0 for Zn and Cd ions and 6.0 for Pb and Cu ions. The adsorption process was found to follow the pseudo second-order kinetic model and Langmuir isotherm model.

The maximum monolayer adsorption capacities of Zn²⁺, Pb²⁺, Cu²⁺, and Cd²⁺ ions onto zeolite at 20 °C were 12.025, 15.96, 12.247, and 16.17 mg.g⁻¹, respectively, showing a significantly greater removal efficiency, which agrees with the previously obtained results. Ion exchange was the principal mechanism controlling the removal of Zn²⁺ and Cd²⁺ from aqueous solution by synthesized zeolite. In contrast, the main mechanism controlling the process of the removal of Pb²⁺ and Cu²⁺ ions is proposed to be either surface adsorption or precipitation. The proposed adsorption mechanism in Figure 27 shows that the adsorption of Zn²⁺, Pb²⁺, Cu²⁺, and Cd²⁺ on zeolite likely can be divided into three major phenomena: precipitation, ion exchange, and surface adsorption [233,234]. The cations exchange is the crucial mechanism for absorbing Zn²⁺ and Cd²⁺ from an aqueous solution. The following equation can represent this process:

(Zeolite-SASR) Na²⁺ + Me²⁺→(Zeolite-SASR) Me²⁺ + Na²⁺

(43)

As a result of industrially applying a synthetic zeolite derived from the SASR be-product, it was demonstrated that the synthesized zeolite effectively removed heavy metal ions from the wastewater sample collected from the Egyptian General Petroleum Corporation (EGPC) (Eastern Desert, Egypt). The targeted levels of heavy metal ion concentrations before and during adsorption using synthesized zeolite are shown in

Table 15. Applying the synthetic zeolites based on SASR-kaolin mixture for removing heavy metal ions from EGPC industrial wastewater sample.

Metals		Cd	Cr	Cu	Mn	Pb	Fe	Ni	Zn	V
$C_0$(mg.L−1)		6.127	7.016	10.294	8.152	11.493	8.219	0.002	17.051	0.01
MLDWHO		0.01	0.01	1	0.1	0.01	1.5	0.1	1	0.002
Number of Cycles	1	U. D	U. D	U. D	2.1	U. D	3.1	U. D	U. D	U. D
	2	U. D	U. D	U. D	0.6	U. D	1.1	U. D	U. D	U. D
	3	U. D	U. D	U. D	0.05	U. D	U. D	U. D	U. D	U. D
	4	U. D	U. D	U. D	U. D	U. D	U. D	U. D	U. D	U. D
	5	0.08	0.05	1.3	2.12	0.51	2.5	0.054	1.66	0.027
	6	0.45	2.01	1.7	7.12	1.07	4.5	0.098	3.87	0.089

U. D: Under detection limit; ${\mathrm{M}\mathrm{L}\mathrm{D}}_{\mathrm{W}\mathrm{H}\mathrm{O}}$: Maximum limits for wastewater disposal according to WHO.

. However, in the fourth adsorption cycle, the residual heavy metal ion concentrations follow WHO guidelines for wastewater disposal in the marine environment with high removal efficiency [235]. When the number of circulations increases (5th and 6th cycles), the amount of heavy metal ions in the effluent also goes up. As a result, when the pollutants are enriched to a specific level, the adsorbent needs to be cleaned to achieve closed recycling. Finally, the synthesized zeolite can be utilized to safeguard industrial wastewater drainage systems.

It was noted that zeolites synthesized based on oil shale ash by an alkaline hydrothermal process were effectively used to extract Pb²⁺ and Cd²⁺ metal ions [236]. Synthesized Na-A zeolite from class F fly ash (FA) and modified oil shale ash (MOSA) by alkaline fusion followed by refluxing was used as an adsorbent for lead, zinc, and chrome [237]. Fly ash hydrothermally modified with NaOH solution was utilized to synthesize zeolites for Cd²⁺ adsorption, which showed effective removal of Cd²⁺ from the wastewater source [238]. Applying a low-temperature roasting process, the novel adsorbent was prepared from fly ash and solid alkali (NaOH). This adsorbent was manufactured under the following conditions: the pristine fly ash and NaOH mass ratio was 5:8, the calcination temperature was from 300 to 350 °C, and the reaction time was 3 h. The obtained results showed a better adsorption capacity of Cd²⁺ on modified fly ash (MFA) [239]. A zeolitic material that was prepared from coal fly ash (CFA) through NaOH fusion treatment, followed by hydrothermal processing, was applied to remove heavy metal ions such as Ni²⁺, Cu²⁺, Cd²⁺, and Pb²⁺ from the wastewater source [240]. The adsorption capacity of the synthesized Na-A zeolite based on the fusion and hydrothermal treatment of oil shale was evaluated by measuring the maximum removal efficiency of Cu²⁺, Ni²⁺, Pb²⁺, and Cd²⁺ from aqueous solutions [241]. A new composite material was formulated based on fly ash (FA), meta-kaolin (MK), and TiO₂ to form a micro-porous zeolitic material with enhanced photocatalytic properties for the adsorption of methylene blue (MB) dye [242]. A modified geopolymer based on fly ash was used to study the effectiveness of adsorbing Cd²⁺ ions from aqueous solutions, as previously described by Javadian [243]. The alkaline fusion-hydrothermal process was used to synthesize zeolites from Brazil oil shale ash. It was noted that the synthesized zeolites are composed of likewise mixed phases (Na-A zeolite, Na-X zeolite, hydroxy sodalite, and quartz) [244].

Yu [245] investigated the efficiency of zeolite-based fly ash to adsorb and remove Ni²⁺ ions from aqueous solutions. This study examined the adsorption capacities of synthetic and commercial zeolite (4A), indicating that zeolite 4A had a lower adsorption capacity than fly ash-based zeolite (75.6 mg/g). Chen [246] developed four different fly ash-based zeolites to adsorb Ni²⁺ ions from aqueous solutions and achieved removal efficiencies ranging from 92.5% to 96.2%. A specific kind of geopolymer made from fly ash showed a significant effect in removing Pb²⁺ ions from aqueous solutions. The maximum removal efficiency of 90.66% has been achieved at 5 pH and a contact time of 2 h [247]. In order to remove Pb²⁺ ions from aqueous solutions, fly ash was employed to create the hydroxysodalite zeolite. The maximum adsorption efficiency of synthetic zeolite was 98.1%, demonstrating the adsorbent’s excellent efficiency [181].

Multi-cation wastewater containing Pb²⁺, Zn²⁺, and Cd²⁺ was treated using a type of zeolite synthesized from fly ash. Pb²⁺ and Zn²⁺ ions were efficiently and selectively removed with 100% and 70% adsorption efficiencies, respectively, compared to Cd²⁺ ions with a 60% adsorption efficiency [248]. Shyam [249] used fly ash adsorbent modified with CaCO₃ to minimize the concentration of Pb²⁺, Ni²⁺, and Cr⁶⁺ ions from aqueous solutions. This study found that 1:10 CaCO₃/FA eliminated Pb²⁺ up to 90%, Ni²⁺ 50%, and Cr⁶⁺ 30% under ideal conditions. In the same way, the adsorption of several heavy metal ions from aqueous solutions was also studied using FA coated with chitosan. It was noted that within 3 h, the produced adsorbent exhibited satisfactory performance. The adsorption capacities of different metal ions have been reported as follows: 36.22, 28.65, 55.52, and 19.10 mg/g for Cr^{3+, 6+}, Cu²⁺, Zn²⁺, and As⁵⁺ ions, respectively [250]. The highest Cd²⁺ ions removal efficiency of 84% was obtained under the best-recommended conditions. The isotherm and kinetic studies indicated that the Langmuir and pseudo second-order models were in good agreement with the adsorption data. Also, they employed FA to synthesize zeolite through the fusion process and showed effective adsorbent to remove Cr⁶⁺ ions from aqueous solutions. More than 80% of the Cr⁶⁺ ions were eliminated, and both the Freundlich and Langmuir isotherm models provided an excellent fit to the equilibrium data. The thermodynamic studies revealed that the process was spontaneous and endothermic, and the adsorption kinetics exhibited pseudo second-order behavior [251]. In conclusion, fly ash has been converted into various ceramic products, zeolite materials, and geopolymers, which in turn are used for wastewater treatment [252].

Table 16. Adsorption conditions of various heavy metals by zeolite-based adsorbents [252].

Heavy Metals and Nutrients	Adsorbent	Removal Efficiency	Ref.
Cd2+	ZFA-600 Zeolite X TiO2/FA) FA-Z MG-Z and MT-Z	84% 100% 80% 60% ~98% and 75%	[243]
			[253]
			[254]
			[248]
			[255]
Co2+	MCM-41	~90%	[256]
Cr3+		~90%
Cr6+	ZFA	>80%	[257]
	CFA-FeOOH	84.9%	[258]
	MSFA/PPy		[259]
	ZFA-Na-A		[260]
	Zeolite X		[261]
	Chitosan/CFA		[262]
Cu2+	FA-MS	98%	[263]
	FA-IOT-Geo	98.3%	[264]
	CFA-Geo	93.9%	[265]
	MPF		[266]
	TiO2/FA	90%	[254]
	MG-Z and MT-Z	100%	[255]
	Ag-Fe3O4/FA		[267]
Fe3+	Zeolite NaeP1	100%	[268]
Hg2+	Zeolite LTA ZFA HMAS zeolite	94% 91.27% ~95%	[269]
			[270]
			[271]
Mn2+	CFA Zeolite	100%	[272]
	Zeolite	100%	[273]
Ni2+	Zeolite		[245]
	FA-Na-P/TEA FA-Na-X FA-Na-P/Na-Br FA-Na-P	96.2% 95.5% 95% 92.5%	[246]
	Zeolite X	95%	[253]
	MG-Z and MT-Z	~50% and 52%	[255]
	PB/FA—SA-FA	100%	[274]
Pb2+	ZCFA Geopolymer FA-Z 1:10 FA MG-Z and MT-Z PB/FA—SA-FA	98.1%	[181]
		90.66%	[247]
		100%	[248]
		>90%	[249]
		100%	[255]
		100%	[274]
Zn2+	FA-Z	70%	[248]
	FAICS		[250]
	FA		[275]
	FA		[276]
NH4+	PB/FA—SA-FA	100%	[274]
	Zeolite NaeP1	~61%	[277]
	Z-P1	65.2%	[278]
PO43+	ZHLO	60%	[279]
	ZFA	91%	[280]
	Z-P1	92.3%	[278]

summarizes the adsorbents synthesized and their metal ion removal efficiency using various preparation methods.

5. Future Recommendations

The synthesizing and application of zeolite from useless by-product waste need further investigation.

To enhance the sustainability and efficiency of salt-roasting processes, alternative methods must be explored for recycling released gasses and barren solutions, along with extracting valuable metals from a secondary vanadium source. Additionally, producing chemical compounds for further use in the relevant process needs further studies.

Other synthetic pathways for producing zeolite materials with high surface area, adjustable molecular dimensions, and adsorption capacities are needed for industrial by-product application. Further research is needed to identify low-cost raw materials for the synthesis of zeolites.

In addition, to reduce the production costs, it is necessary to develop accessible synthetic routes. Kinetic studies and energy calculations are needed to investigate how nucleation occurs during zeolite synthesis. Further studies on the regeneration of synthetic zeolite are required to increase the cost-effectiveness of the process.

6. Conclusions

The shortage of vanadium primary resources from high-grade ores makes it necessary to find a suitable alternative that covers this shortage. Also, the world vision concerning the solution to climate change and environmental pollution problems is based on decreasing or completely getting rid of the amounts of the produced waste. One of these promising alternatives is the utilization of different wastes containing vanadium to decrease their environmental adverse effects and, at the same time, enhance their economic value by converting them into valuable products.

The current study aims to present and analyze the worldwide available alternative vanadium-bearing waste resources, proprieties, reserves, and processing technologies. The heavy oil fly ash “HOFA” and vanadium slag showed a promising alternative for the production of vanadium metal due to their high content of vanadium and the cost-effectiveness of their processing to extract vanadium metal compared to processing their corresponding low-grade ores.

Vanadium can be extracted from these wastes using pyro- or hydro-metallurgical methods or a combination. The hydro-metallurgical extraction can be carried out using one of the vital mineral acids such as HCl, H₂SO₄, or HNO₃, and previous studies reported that HCl and H₂SO₄ have comparable leaching efficiencies. H₂SO₄ is preferred due to its significantly lower cost and its wide application in the leaching of many metals. In addition to V and Ni, H₂SO₄ effectively dissolves almost all metals in HOFA and other vanadium-bearing wastes, such as Mo, Mg, Mn, and others. In addition to acidic leaching reagents, alkaline leaching reagents like NaOH, Na₂CO₃, and NH₄OH can be used for leaching vanadium from their containing wastes. The previous study showed that NaOH has the highest selectivity and leaching efficiency for V compared to other alkaline leaching reagents.

Two-stage leaching can be used for more selectivity and productivity of vanadium from their containing wastes. The first stage includes the alkaline leaching of high-contented and alkaline leachable vanadium using NaOH, followed by the second acidic leaching of acidic leachable elements using H₂SO₄. A combination of pyro- and hydro-metallurgical processes is carried out for more effective extraction of vanadium from their containing wastes. Firstly, pyro-metallurgical processes like roasting are carried out to convert low-leachable vanadium to high-leachable phase. Alkaline salt roasting using NaCl and Na₂CO₃ converts contained vanadium into water-leachable phase NaVO₃. Previous studies showed that the roasting temperature is the main controlling parameter on the efficiency of vanadium extraction. The hydro-metallurgical processes include leaching the modified (roasted) wastes using water or a very low-concentration acidic leaching process. Recent studies confirm the cost-effectiveness of the salt roasting process followed by the water leaching process for treating vanadium-bearing wastes, which showed high percent recovery and purity of the extracted vanadium. After leaching, V can recover from the leaching solution by chemical precipitation, solvent extraction, or ion exchange processes.

Complex processing of vanadium-bearing wastes, including firstly the extraction of valuable metals followed by the complete utilization of the resulting residue in the production of adsorbent, i.e., zeolite with different composition and properties to be used in wastewater treatment, is very promising and applied alternative for complex processing of these wastes.