Ecological treatment of El Kriymat boiler ash for recovering vanadium, nickel and zinc from sulfate leach liquor

The primary goal of this work is to develop a technology that allows for the recovery of metal values from waste products, thereby promoting the wise and efficient use of our nation's resources. To achieve this goal, an industrial waste of El Kriymat boiler fly Ash was used for recovering its content of vanadium, nickel and zinc. About 97, 95 and 99% respectively of these economic elements were first dissolved from boiler fly ash magnetic concentrate (after physical concentration). Leaching experiments using optimum conditions include: 180 g/L sulfuric acid concentration and 4% solid/solid proportion manganese dioxide acts as an oxidant at 80 °C. The recovery of vanadium (V) metal ions was carried out using 3% Alamine 336 in kerosene at an equilibrium pH value of 0.9. Subsequently, 15% sodium sulfide solution was used for co-precipitation of nickel and zinc metal ions in the raffinate solution at pH value of 3.5.


Introduction
Vanadium, nickel and zinc are widely used in aerospace, stainless steel, metallic alloys, aircraft industries, electronic information, battery, pigments, ceramic industries, industrial production and many other technical fields for their unique properties of corrosion resistance, high temperature resistance and wear resistance, which have greatly stimulated the demand in recent years [1,2]. Naturally, vanadium, nickel and zinc are present in many ores, but their recovery from the natural resources is not sufficient for their uses. Therefore, the world is interested by the ashes as additional and economic resources of these metal values [3][4][5][6]. Internationally, massive volumes of ash are produced by Egypt's oil-fired power plants; roughly 4000 metric tons of oil ashes are produced annually [7].
The US EPA classifies fly ash created by power plants (which amount to several million tons per year) as a particular waste, but minor portions of these ashes are re-used [8,9]. Due to its varied assembly and physical qualities, determining the nature of fly ash becomes more difficult. The constitution of fly ash varies based on the type of coal used, boiler used, operating conditions, post-combustion technology used, and the power plant technology used [10,11]. Boiler ash is a by-product of burning of fuel that contains high amounts of V, Ni, Zn, and other precious elements that water can be used to remove it, although alkaline or acid solutions work better [6,[12][13][14][15]. The presence and concentration of various heavy metals in fly ash residue can be defined by feedstock traits, operating conditions and flue gas treatment facilities [16]. It is suggested that fly ash can be and is currently used as a various construction material [17][18][19][20], thermochemical energy storage (TCES) material, a H 2 S removal agent [21], a low-cost adsorbent for gas and water treatment [6], and in some instances, it can be converted to a zeolite-based catalyst [22][23][24].
Heavy metals in ashes posed varying levels of environmental contamination danger from both an environmental and a use standpoint. According to the oil origin and ash consumption conditions, V content in oil ashes ranges from 2.5 to 30%. Since 1960, the recovery of V from oil ashes has attracted attention not only as a matter of economic concern (metals recovery), but also as a matter of environmental concern (people, waterways, and soils) [25][26][27]. A method has been devised for extracting V, Ni, and other elements from fly and boiler ash while also rendering the ash safe [28]. Elomaa et al. [29] investigate the behavior of fly ash during a 24-h leaching process in different concentrations of water, ethaline, and various mineral and organic acids. The types and concentrations of the reagents used as well as the heavy metal itself affected the leaching efficiencies. The selective leaching of precious metals, however, focused on Fe. A sustainable leaching method needed to be found to lay the groundwork for future work on metal recovery from fly ash. The highest Zn extraction (> 80%) was achieved using 3 M H 2 SO 4 and 7 M HCl, which also dissolved Cu (> 87%) and Ni (> 65%). The use of HCl allowed for the efficient leaching of Pb; full extraction was provided by 3 and 5 M HCl, whereas H 2 SO 4 was unable to extract Pb due to the formation of PbSO 4 ; similarly, Pb precipitated as PbC 2 O 4 in oxalic acid. Ethaline could extract 50% of Pb and had a high selectivity toward Fe.
Generally, leaching or extraction, thermal treatment and solidification/chemical stabilization are considered as common approaches for fly ash treatment [6,16]. The most common methods for extracting vanadium and other metals from ashes are hydrometallurgy (acidic, alkaline) and pyro-hydrometallurgy (roasting) [7,[30][31][32][33][34][35][36]. In this case, a temperature of 200 °C and a partial oxygen pressure of 15 atm are used. Amer [37], treated Egyptian boiler ash with 60 g/L H 2 SO 4 to recover both of V and Ni. The aforementioned procedure necessitates high temperatures and pressures, which are not economically feasible on an industrial scale. Consequently, Tokuyama et al. [38] recovered 80% of V and Ni using C467 and CR20 respectively ion exchange resins, Guibal et al. [12] reported that, several amine extractants (Primene JM-T, Amberlite LA-2, Alamine 336, and Alamine 304), a quaternary ammonium salt (Aliquat 336), and a chitosan sorption technique (for low-metal concentration solutions) were used to recover V and associated important elements. In acidic and alkaline media, Alamine 336 in kerosene is an excellent extractant for numerous metal ions [39][40][41][42]. Vahidi et al. [5], used D 2 EHPA, Cyanex 272, and their combinations in various quantities to remove Ni and V from sulfate leach liquors of power plant fly ash.
Precipitation of metals as metal sulfides, on the other hand, is a feasible approach to recover metals from mining solutions. Because many metals are only sparingly soluble as sulfides, sulfide precipitation is advantageous. It's a method for achieving extremely low residual metal concentrations (less than 0.01 mg/L). Sulfide precipitation's final concentration is proportional to pH, and the process is far more efficient than hydroxide precipitation [43]. Tokuda et al. [44], claimed to have controlled sulfide precipitation at pH 1.5 for CuS, 4.5 for ZnS, and 6.5-7.0 for NiS precipitation. It was discovered that an amount of H 2 S equimolar to a specific metal was sufficient to precipitate the metal virtually completely. Tokuda et al. and Sampaio et al. [44,45] explored the selective removal of Zn from a combination of Zn and Ni with Na 2 S. At pH 5, the selectivity was enhanced to 99% by reducing super saturation at the dosing points by lowering influent concentrations. Sphalerite formed from Zn, while Ni-based amorphous particles with a pH-dependent stoichiometry.
The primary goal of this study is to find the best environment and approach for extracting V, Ni, and Zn from boiler ash utilizing a sulfate solution, followed by solvent extraction for vanadium over the related components and direct precipitation of Ni and Zn from the resulting liquor solution. The commercial extractant (Alamine 336) in kerosene was successfully used in this investigation to separate and recover V from the prepared sulfate solution. Solvent extraction has been identified as one of the finest ways for extracting and separating metals in terms of cost, environmental concerns, and achieving high separation and extraction levels.

Materials and method
Almost all of the chemicals used in the experiments are analytical grade, such as sulfuric acid, ammonium hydroxide, sodium chlorate, and kerosene. Alamine 336 [Sigma Aldrich], Octanol [Adwic], Na 2 S [Panreac (PRs)]. Before conducting leaching and recovery studies upon El Kriymat boiler ash concentrate, different samples were first collected from El Kriymat electric station, good mixed and quartered to get the representative working sample (1 kg). The latter was then subjected to physical beneficiation using High Tension Magnetic Separation; a Carpco Inc. Florida USA Model MLH.13.111-5 Left Magnet at 1.0 Amp. The obtained magnetic fraction was first analyzed using Axios advanced WDXRF-PANalytical (Netherland) for its major and trace elements. In this regard, an atomic absorption spectrometer (Unicame 969) was utilized for quantitative analysis of V, Fe, Ni, and Zn at proper wavelength during all experimental work (leaching, extraction and precipitation) [46]. Finally, the products of the metals of interest namely; V, Ni and Zn were confirmed using both scanning electron microscope (ESEM-EXL30 Philips type) coupled with an energy-dispersive X-ray analyzer (EDX unit system) and (PHILIPS PW 223/30) X-ray diffraction technique (XRD). The latter was linked to a diffractometer with an automatic sample changer, Device model Malvern P analytical Empyrean (2020) (Netherlands) (21 positions).

Processing procedure
First, the leaching investigation and comprehensive chemical characterizations were carried out on the magnetic fraction at 1.0 Amp. Before executing V, Ni, Fe, and Zn leaching from petroleum ash sample tests, the theoretical thermodynamic factors were estimated using the computer program HSC chemistry 5.1(the world's favorite thermochemical software it is designed for various kinds of chemical reactions). The ∆G in reactions of the concerned element oxides with the acids HCl, HNO 3 , and H 2 SO 4 were studied to determine the optimal acid for leaching and the appropriate leaching temperature. Some experiment sets were carried out to investigate the parameters affecting the leaching efficiencies of the concerned components in the hydrometallurgical processing to the magnetic fraction of petroleum ash samples, such as temperature and leaching time. The study of the effect of reaction temperature on the rate of leaching of the elements in question is also part of the leaching kinetics. The concentrations of these elements in the leachates solutions after filtration were determined according to the following equation:

Preparation of pregnant solution
The workable sulfate leach liquor for V, Ni, and Zn extraction was made by combining 645 g of the magnetic fraction concentrate from boiler fly ash with 180 g/L H 2 SO 4 under optimal agitation leaching conditions. The final volume of the produced sulfate leach liquid (after filtering and washing with distilled water) reached 6.45 L per 645 g boiler ash magnetic concentration, and its pH value and concentrations of metals of interest were determined.

Separation of vanadium and iron using alamine 336
In this case, a solvent extraction procedure using Alamine 336 in kerosene with octanol as a modifier was used to extract V and Fe under ideal conditions. At room temperature (30 ± 5 °C) and pH (0.6-1.5), solvent extraction tests were carried out by mixing equal quantities of aqueous solution and Alamine 336 (10 mL each) in a separating funnel [organic/ aqueous (O/A) volumetric ratio of unity]. For varying amounts of time, both stages were contacted (1-7 min). After equilibration, phase separation was performed, and the extraction efficiencies of V, Fe, Ni, and Zn in the raffinate solution were calculated. To improve V stripping process, the type and concentration of the stripping solution, shaking time, and phase ratios [aqueous/organic (A/O)] were investigated. In the presence of NaClO 3 as an oxidant, the V strip solution was exposed to direct V precipitation as V 2 O 5 . xH 2 O. The hydrated vanadium oxide precipitate was washed with distilled water, dried at 120 °C for an appropriate time period, and the product's purity was chemically analyzed and then confirmed using XRD instrumental equipment.

Direct precipitation of nickel and zinc
The raffinate solution, which was nearly devoid of V, but still contained some Fe metal ions, was then used to separate Ni and Zn using Na 2 S solution. To manage the precipitation process, relevant effective precipitation parameters, such as Na 2 S concentration, pH values, and precipitation time at room temperature (30 ± 5 °C), were investigated. (1) Leaching efficiency (%) = Element conc. in the leach liquor Element conc. in the original ore × 100

Chemical composition of the working boiler ash
The physical beneficiations of petroleum ash at magnetic fraction of 1.0 Amp produce about 65% of the bulk original sample, and its chemical composition was summarized in (Table 1). The studied magnetic fraction is mainly composed of high concentration of 18 The working sample, on the other hand, contains high amounts of valuable and strategic nuclear elements like U, Th, Cd, Co, Mo, Zr, and Hf. The presence of these nuclear elements has been linked to organic matter found in fissures and fractures as a result of oil seepage following tectonic processes. [47].

Results of theoretical aspects of thermodynamic
Using the HSC chemistry 5.1 computer programs, the free energy for the reactions of various oxides leaching in HCl, H 2 SO 4 , and HNO 3 at 50 °C was calculated. From the obtained data summarized in (Fig. 1A), the ∆G for all oxides is high negative, it is reasonable to expect that H 2 SO 4 will be more effective than HCl and HNO 3 . Moreover, calculating the free energy of the concerned elements oxides with H 2 SO 4 at different temperatures, the results of −∆G (Fig. 1,  B) showed that the free energy of vanadium oxides (V 2 O 5 and V 2 O 3 ) dissolution with sulfuric acid is expected to be the most easily soluble, whereas the free energy of the elements oxides varies in the sequences of V 2 O 5 , V 2 O 3 , NiO, and ZnO ( Fig. 1, B). The dissolution reactions of different oxides with H 2 SO 4 were illustrated as follows:

Sulfuric acid agitation leaching
In terms of hydrometallurgical processing of separated magnetic fraction from petroleum ash sample, the effects of temperature and agitation time upon the concerned elements were performed. The previously published optimum condition (Abd El-Hamid and Abu Khoziem), [48], was used to further investigate the leaching of fly ash minerals using sulfuric acid. The optimal conditions were typically 180 g/L H 2 SO 4 acid, 4% (solid/solid ratio) MnO 2 as an oxidant agent, a solid/liquid (S/L) ratio 1/10, and -200 mesh grain sizes at different temperature and agitation time.
In fact, the effect of temperature on the leaching efficiency of the concerned elements was explored using a series of tests at room temperature, 50, 60, 70, and 80 °C, while keeping other variables constant: 180 g/L H 2 SO 4 and S/L ratio of 1/10. According to the data acquired, increasing the temperature enhanced the leaching efficiency (Fig. 1C). As a result, higher temperatures accelerate molecular motion and increase the number of atoms colliding with H + , demonstrating that dissolution is controlled by a chemical reaction. These results are consistent with the published data, [49,50]. Although increasing the temperature more than 80 °C improves leaching efficiency, the authors do not recommend it because of the high consumption of acid caused by the higher rate of water evaporation and the difficulties of filtration. As a result, 80 °C was chosen as the optimal temperature for energy conservation that the leaching efficiencies at that temperature reach 76, 57.5 and 80.4% for V, Ni, and Zn, respectively. This is because the energy available for atomic and molecular collisions increases as the leaching temperature rises [51].
On the other hand, another series of leaching tests was carried out to determine the effect of agitation time on leaching efficiency of the concerned components, this time increasing the agitation period from 4 to 10 h while keeping the other variables unchanged. The obtained data (Fig. 1D) revealed that, as the agitation time was prolonged up to 8 h, the leaching efficiency of V, Ni, and Zn improved, reaching maximum values of 96.5, 94.8, and 99.1%, respectively. After 10 h of agitation, the leaching efficiency began to decline. This is due to VOSO 4 coats the unreacted particle cores' surfaces. As a result, the leaching process begins to slow [52].

Agitation leaching kinetics study
The effect of reaction temperature on the leaching rate of the relevant elements is discussed here to provide some insight into the kinetics of leaching for V, Ni, and Zn. At temperatures ranging from 30 to 80 °C, with 200 mesh particle size, 180 g/L H 2 SO 4 acid concentration, 4% (w/w) NaClO 3 as an oxidizing agent, and a 1/10 solid/liquid ratio, this experiment is carried out at intervals of 2-8 h. In a recent article, the kinetics of V leaching were investigated in depth [52], and here it will be studied for the elements Ni and Zn. The dissolution of these elements from concentrated boiler ash happens as a liquid-solid reaction at the interface of the two phases in the presence of H 2 SO 4 acid and NaClO 3 as oxidizing agent. The unreacted or shrinking core model is the most often used model for describing fluid-solid reactions (SCM), [53]. The rate-controlling step is the slowest. The experimental results in Fig. 2A and B are connected to various kinetic models for solid-liquid processes to get the kinetic equations for the dissolution of Ni and Zn as examples of the concerned elements as the following: where K is the apparent reaction rate constant (min −1 ), t is the leaching time and X is the reacted fraction which is expressed as X = % metal leaching/100.
The influence of temperature on Ni and Zn leaching from concentrated petroleum ash ( Fig. 2A and B) shows that as the temperature rises, the leaching rate of the two elements increases. In the process of leaching petroleum ash, there are two stages. The percentage of removed Ni and Zn increased dramatically over the first 8 h, although these elements leaching efficiencies fell significantly with time. This conclusion is consistent with the findings in Fig. 1D as well as the published data [52,54].
( Figure 2C and D) shows the relationship between 1−(1−X) 1/3 and the Ni and Zn leaching time at various temperatures respectively. Although the chemical controlled model is considered the most appropriate for the description of Ni and Zn dissolution kinetics based on the correlation coefficient (R 2 ), while the relationship between 1-2/3(X)-(1-X) 2/3 and the concerned elements leaching time at various temperatures ( Fig. 2E and F) for Zn and Ni respectively shows that, the mixed model describes the kinetics of Zn and Ni dissolution from petroleum ash sulfuric acid solution. The probable explanation for the chemical and mixed reaction control mechanism under the specified leaching conditioning may be attributed to the chemical form in which the Ni-and Znbearing particles exist in the petroleum ash, which also is consistent with the previously published data [52,55].

Preparation of pregnant solution
The pregnant leach liquor was prepared using the identified optimal sulfuric acid agitation conditions for future recovery of V and other associated economic elements, such as Ni and Zn. The applied optimal leaching conditions were typically 180 g/L H 2 SO 4 acid, 4% (solid/solid ratio) MnO 2 as an oxidant agent, 1/10 (S/L) ratio, and − 200 mesh grain sizes at 80 °C temperature and 8 h agitation time. As a result, 645 g of the boiler ash magnetic concentration was utilized to make 6.45 L of sulfate leach liquor. The chemical analysis of the prepared sulfate leach liquor of the working boiler ash concentrate is summarized in (Table 2), and its pH value is 0.31.

Recovery of vanadium using alamine 336
According to Abd El-Hamid et al. [56], U, Hf and Zr were first recovered from petroleum fly ash sulfate leach liquor using Dowex 1X8 as strong basic anion exchange resin technique. For the concerned elements (Hf, Zr, and U), the ideal adsorption conditions are 5 ml/min flow rate at pH 1.5, with a reducing agent applied to avoid Fe and V adsorption. After oxidizing the solution, a solvent extraction procedure using Alamine 336 in kerosene and octanol (as modifier) was utilized to separate vanadium from the raffinate of the produced sulfate leach liquor. The expected reaction mechanism of the amine-based extractants occurs in two steps, protonation and ion exchange, as shown below: where [R 3 N] represents Alamine 336. Several parameters that govern the extraction process were investigated as follows: This factor was considered as the most influenced parameter that encourages V extraction and separation away from the other metal ions e.g., Ni and Zn from the prepared sulfate solution using Alamine 336 in kerosene. The data in Fig. 3A reflected the effect of increasing pH of the produced sulfate solution from 0.6 to 1.5 on the extraction efficiencies of V, Fe, Ni, and Zn. This factor was investigated using 2% Alamine 336 in kerosene with a 1/1 O/A volume ratio and a contact period of 3 min. The V extraction efficiency was found to be 80.1% at pH 0.9 and steadily fell to 67.3% at pH 1.5. This was due to the partial hydrolysis and precipitation of V at lower pH value to obtain water-insoluble vanadium compounds Na 2 V 12 O 31 [57]. The loading efficiencies of Fe, Ni, and Zn, on the other hand, gradually increased as the pH of the solution increased. Thus, Alamine 336 shows a greater affinity for V over Ni and Zn at a lower solvent concentration and lower pH values; as V form anionic species in acidic solutions. Alamine 336 is capable of forming oil soluble salts of anionic species at low pH. It contains a basic nitrogen atom, typically can react with a variety of inorganic and organic acids to form amine salts, which are capable of undergoing ion exchange reactions with V anions [58]. As a result, the pH value of 0.9 is deemed ideal.
Alamine336 concentrations ranging from 2 to 4% were tested at equilibrium pH 0.9 with an O/A volume ratio of 1/1 and a contact time of 3 min. As the Alamine 336 concentration grew from 2 to 3%, the V and interfering Fe extraction efficiencies improved from 80.1 and 67.2 to 97.9 and 79.5%, respectively, according to the obtained results (Fig. 3B). Otherwise, the extraction efficiencies of Ni and Zn at this concentration were only 6.1 and 8.9%, respectively. When the solvent content was increased to 4%, all of the components of interest showed a limit increase.
Contacting 3% (v/v) Alamine 336 with a sulfate leach liquid of pH 0.9 at varied contact times ranging from 1 to 7 min at a 1/1 O/A ratio was used to study the influence of contact time. The acquired results (Fig. 3C) revealed that during a contact period of 3 min, V's extraction efficiency reached its maximum value (97.9%). In terms of Fe extraction efficiency, it was discovered to be related to contact time.
This impact was investigated using 3% Alamine 336, a contact period of 3 min, and a pH of 0.9 at varying O/A ratios of 2/1…0.1/1… 1/2. The data in (Fig. 3D) showed that, while V extraction efficiency at O/A ratios of 2/1 (100%) was higher than O/A ratios of 1/1, V extraction efficiency at O/A ratios of 1/1 was lower (97.9%). However, due to dilution, this ratio was not used because it results in lower V concentration in the organic phase. The V extraction efficiencies at O/A ratios 1/2 and 1/3, on the other hand, were 60.1 and 45.8%, respectively, which were lower than 1/1 (97.9%). In fact, the best O/A ratio for maximum V extraction efficiency was 1/1. By constructing a McCabe-Thiele diagram (Fig. 3E), it is easily concluded that, only two experiment stages are required to achieve a complete extraction of vanadium from the leach liquor when applying a mixer settler.

Stripping process
The stripping process was carried out to achieve maximum V stripping efficiency and, on the other hand, to separate it from the co-extracted Fe. According to Hart, and Hoogerstraete et al. [59,60], Fe could only be stripped by forming a water-soluble Fe complex with ethylene-diaminetetra-acetic acid (EDTA). The latter forms a stable, highly water-soluble chelate complex with Fe (III) with stability constant (EDTA-Fe 3+ complex of 25.1). For this purpose, a solution of 0.5 mol/L EDTA solutions was first used for Fe stripping from the loaded Alamine 336. Almost complete stripping of Fe free from V from the organic solvent was carried out at the optimum conditions of 0.05 mol/L EDTA solution, contact time 5 min at A/O volume ratio of 1/1. On the other hand, the following stripping parameters would be investigated for optimizing the stripping of V from the organic phase.
Different concentrations of H 2 SO 4 , HCl, HNO 3 , Na 2 CO 3 , (NH 4 ) 2 CO 3 , and NaCl were investigated at A/O ratio of 1/1 and a contact time of 3 min, 0.5 mol/L. When compared to the other stripping agent types, the data in (Fig. 4A) demonstrated that 0.5 mol/L H 2 SO 4 acid had the highest stripping effectiveness of 21.5% after separation and determination of V in the strip solution.
This effect was achieved by contacting the laden solvent with varied concentrations of H 2 SO 4 acid solution ranging from 0.5 to 3 mol/L at a 1/1 A/O volume ratio and a contact time of 3 min. Raising the H 2 SO 4 concentration from 0.5 to 2.5 mol/L resulted in a significant increase in V stripping efficiency from 21.5 to 80.8% in (Fig. 4B), while increasing the strip solution concentration to 3 mol/L resulted in a relatively small improvement in V stripping efficiency to 82.9%. To regenerate V from the loaded solvent, 2.5 mol/L H 2 SO 4 was already chosen. When 2.5 mol/L H 2 SO 4 solutions were contacted with the loaded Alamine 336 at an A/O ratio of 1/1, (Fig. 4C) depicted the effect of varied contact time from 3 to 15 min. At a contact time of 10 min, it was discovered that 94.7% of extracted V was stripped from the organic solvent.  5 0.6 0.7 0.8 0.9 1.0 1.1 1.2 1.3 1.4 1.5 1.6  However, increasing the time by more than 10 min had no discernible influence on V stripping efficiency. Using a 2.5 mol/L H 2 SO 4 solution and a 10-min stripping duration, the equilibrium state of V stripping process from the loaded solvent was examined using varied A/O ratios of 2/1, 1/1, 1/2, and 1/3. (Figure 4D) showed that the loaded V was stripped completely (100%) at an A/O ratio of 2/1, compared to 94.7%, 62.9%, and 49.5% at A/O ratios of 1/1, 1/2, and 1/3, respectively. To counteract the diluted V in the strip solution, the ideal applied A/O volume ratio was 1/1. According to the findings of this study, a 2.5 mol/L H 2 SO 4 acid concentration should be used to remove 94.7% of V with a contact time of 10 min and an A/O ratio of 1/1. By generating a McCabe-Thiele diagram (Fig. 4E), we can simply determine that just two experiment stages are required to achieve total vanadium removal from the loaded solvent when using a mixer settler.

Preparation of vanadium oxide red cake
The final vanadium product was made using a V-rich strip sulfate solution that assayed 5.84 g/L of vanadium. The vanadium-rich solution, on the other hand, was adjusted to pH 2.5 using NaOH and subsequently oxidized with NaClO 3 . At 75 °C, almost full precipitation of vanadium (99.2%) as a crimson cake was achieved after 2 h of stirring. To eliminate the related salts, the final product was filtered and rinsed with distilled water. The obtained red cake was then dried in the oven at 120 °C for a suitable period of time. XRD analysis was used to identify the latter, as shown in (Fig. 5). The purity of the final product (hydrated vanadium oxide) was 95.9% with linked sodium ions, according to the chemical analysis.

Direct precipitation of nickel and zinc products
After almost complete removal of V and about 79% of Fe from the prepared pregnant sulfate leach liquor, the raffinate solution (containing 5.07 g/L Ni and 3.83 g/L Zn) was then subjected to nickel and zinc sulfide precipitation using Na 2 S solution. Metal sulfide precipitation is an effective method for removing heavy metal ions (Ni, Zn). The main advantages of using sulfides are-the lower solubility of metal compounds than hydroxide precipitates, are not amphoteric, so it can achieve a high degree of metal removal in a shorter time over a wide pH range, better settling properties; it is easier to thicken and de-water compared with hydroxide. The precipitation of Ni and Zn using Na 2 S from the raffinate sulfate solution can be summarized as follows: where M: Zn and/ or Ni. The relevant effective precipitation parameters at room (30 ± 5) temperature are: The influence of pH on Ni and Zn precipitation efficiency was investigated using a 5% Na 2 S solution, 30 min precipitation time at pH values ranging from 1 to 4. (Figure 6A), it's clearly evident that the precipitation efficiency of Ni and Zn was increased from 5.7 and 1.1% to 78.6 and 88.1%, respectively, as the pH value increased from 1 to 3.5. Furthermore, the precipitation efficiencies of both of Ni and Zn recorded 83.4 and 93.7%, respectively, with increasing the associated impurities especially Fe at pH value 4. Indeed, pH 3.5 was already chosen. The raffinate solution was stirred with different Na 2 S concentrations (in the range of 5-20%) for 30 min at final pH value 3.5. ( Figure 6B) showed that, the precipitation efficiencies of both of Ni and Zn reach their maximum values of 98.9 and 99.8% respectively when using 15% Na 2 S concentration. The higher sulfide concentration in the obtained product of Ni-, Zn-sulfide cake was due to an increased volumetric sulfate-reducing activity upon higher sulfate dosing, resulting in more sulfide being produced.
A number of precipitation experiments were carried out for a time ranging from 10 to 30 min to investigate the effect of precipitation time on the precipitation efficiencies of both Ni and Zn from its raffinate solution. The other factors were kept constant at final pH 3.5 and 15% Na 2 S concentration. Although the precipitation efficiencies of both of Ni and Zn recorded 90.3 and 93.1%, respectively at precipitation time 10 min. However, the precipitation efficiencies reached maximum values (98.6% for Ni and 99.8% for Zn) by stirring time for 30 min, (Fig. 6C). Finally, from the previous co-precipitation data, it is worthy to mention herein that, the optimum conditions for co-precipitation of Ni and Zn from the raffinate solution were 15% Na 2 S concentration, final pH value 3.5 and precipitation time 30 min. The final product of Zn and Ni sulfide cake after washing and drying was then ignited to its oxide at 850 °C for 2 h. The chemical analysis of Ni-Zn oxide cake showed purity of 97.1% with associated impurities of 2.3% Ca, 0.5% Co and 0.3% Fe while Ni and Zn represented 54.9 and 42.2%, respectively. The ESEM-EDX area analysis of Ni and Zn precipitate was illustrated in (Fig. 7). Finally, a mass balanced technical flowsheet shown in (Fig. 8), was proposed on the assumption of chemical treating El Kriymat boiler ash concentrate.

Conclusion
• Petroleum fly ash is regarded as a double blessing material.
On the one hand, it is presented as an environmental issue.
On the other hand, it would be a true treasure because it contains valuable elements, so this study aims to try to convert a major environmental problem into a major economic treasure. • Physical beneficiation was first carried out upon the working petroleum ash raw material using High as oxidant was applied upon the concentrate sample for dissolving its main contents of V, Zn, Ni, Cr and Fe which save about of 40% of consumed chemicals from economic and ecological points of view. • Separation of hafnium, zirconium, and uranium was carried out as the first step upon the leach liquor, then selective extraction of V with efficiency of about 98% associated with 77.8% of Fe was carried out using 3% Alamine 336 in kerosene leaving behind both of Zn and Ni in raffinate solution. • For separation of Fe above V from the organic solvent, a 0.05 mol/L EDTA solution was utilized with a contact time of 5 min and an A/O volume ratio of 1/1. • Highly pure hydrated V 2 O 5 was prepared from the strip sulfate solution at pH 2.5 with stirring time 2 h at 75 °C while, Ni-Zn cake was then prepared from the raffinate solution which will be subjected for further treatment for individually separating its contents. • Finally, a technical flowsheet was proposed on the assumption of chemical treatment of El Kriymat boiler ash concentrate. • The current study succeeded to demonstrate a low-cost, echo-friendly hydrometallurgical technique for recovering hafnium, zirconium, uranium, vanadium, nickel, iron, and manganese from petroleum leach liquor. Acknowledgements The authors would like to thank the Nuclear Materials Authority, for financial support during this research. They are also grateful to the members of the Nuclear Materials Authority's Labs for their kind assistance, valuable counsel, and honest assistance throughout the work's analytical control.
Funding Open access funding provided by The Science, Technology & Innovation Funding Authority (STDF) in cooperation with The Egyptian Knowledge Bank (EKB). Ministry of Scientific Research, Egypt Data Availability The datasets generated during and/or analysed during the current study are available from the corresponding author on reasonable request.
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