Techno-economic study on reagents consumption during uranium leaching, a case study: Gattar pilot plant, Egypt

Abstract As a techno–economic study, column and vat leaching experiments of low-grade uranium mineralization were implemented to obtain a deep understanding of reagent consumption during the uranium leaching stage. The leaching index (β) is a techno-economical parameter that must be considered during the decision-making stage of project development in order to arrive at an optimal point based on solution consumption and recovery. The results indicate that 21.2 and 32 g of acid are consumed using an acid concentration of 20 and 40 g/L, respectively, to have a 1% increase in the uranium recovery. Also, 1.0 and 1.8 L solutions using 0.1:1 and 0.2:1 sprinkling intensities are consumed, respectively, to have a 1% increase in the uranium recovery during column leaching period. The results of the leaching index (β) of vat leaching indicate that 4.1 m3 solution and 88 kg acid are consumed to have a 1% increase in the uranium recovery after 90 day. The overall consumed acid after this period was 17.6 kg/ton of ore.


PUBLIC INTEREST STATEMENT
Through this manuscript, a techno-economic study was done through lab (column) and field (vat) leaching experiments of low-grade uranium mineralization to obtain a deep understanding of reagent consumption and recovery during the uranium leaching stage. The leaching index (β) must be considered in order to arrive at an optimal point based on solution consumption and recovery. The optimum conditions obtained from the column experiments were applied to the vat leaching experiments, and the results of vat leaching indicate that 4.1 m 3 solution and 88 kg acid are consumed to have a 1% increase in the uranium recovery after 90 days leaching period. At the same time, the overall consumed acid after this period was 17.6 kg/ton of ore.
In the future, these parameters can be used to process the uranium mineralization of the Gattar area through the heap leaching at the lowest possible economic cost.

Introduction
Mining and extraction processing are the principal sources of national income in many developing countries with vast reserves of profitable minerals; heap leaching technique is becoming more wide spread. As a result, finding cost-effective methods to harvest these minerals, notably uranium, became critical. Because of its low starting cost and technological simplicity, heap leaching is a viable option. For example, Peru, Mexico, and Zambia are in this situation (Van Staden, 2011;Vladimir, 2015). Several countries have turned to heap leaching technology for low-grade uranium mineralization due to the decline of high-grade uranium mineralization (Carlsson & Büchel, 2005;Scheffel, 2010;Shakir et al., 1992;Taylor, 2013). Taylor (2009) describes how uranium heap leaching is becoming more common, with plans to use alkaline leaching on low-grade calcium carbonate-rich materials in the future (Taylor, 2009).
Mineralogy of the ore and gangue is important to metallurgical behavior and, as a result, to an ore's technical and economic viability. The needs for acid curing and agglomeration prior to heap leaching, as well as the acid consumption, metal extraction, and Pregnant Leach Solution (PLS) chemistry during leaching, can be significantly influenced by relatively small variations in gangue mineralogy (Van Staden, 2008).
Certain gangue minerals, which including calcite, are acid-reactive and will dissolve totally in even very dilute sulfuric acid solutions over time. Acidic heap leaching is plainly undesirable for ores that contain a considerable amount of such minerals. The silicate minerals only partially react with sulfuric acid, and the intensity of the acid-gangue interaction varies depending on the acid strength in the leach liquor. These gangue minerals' moderate acid reactivity limits their acid consumption to the point where acid heap leaching of the ore can be commercially viable even with low precious metal grade. By adjusting the acid curing parameters and the acidity of the irrigation fluid in a matrix of metallurgical tests, it allows for a margin of control over acid consumption and the search for an economic optimum between acid cost and leach kinetics (Van Staden et al., 2009). The flow sheet for the leaching stage is very simple, in which solutions of high uranium concentration suitable for the ion exchange extraction process are directed to PLS tanks, and liquids of low concentration of uranium are directed to (ILS) tanks for recycling over the piles (Robertson et al., 2012). The solutions (effluents) from the ion exchange (adsorption) extraction process are recycled with fresh solutions on the surface of the ore pile. An intermittent liquid addition technique is preferred to improve the efficiency of uranium dissolution by leaching (Sheikhzadeh et al., 2005). Continuous addition is punctuated by a pause, accompanied by the addition of a new carrier solution or liquid in periodic leaching. According to several studies on mineral and soil treatment (Cote et al., 2000;Ghorbani, Franzidis et al., 2016;Saririchi et al., 2012), this contributes to an improvement in leaching quality. Intermittent leaching is appropriate for large particle size leaching, and a long pause and dry period are expected to result in more effective leaching (Cote et al., 2000). Furthermore, as opposed to continuous liquid addition operations in industrial leaching scale, the cost of intermittent leaching operations can be significantly lower (Kappes, 2002).
Extensive laboratory studies have been carried out on uranium dissolution from different uranium ore mineralization resources from the Eastern Desert of Egypt using sulfuric acid vial laboratory tests. When optimum conditions are applied in the presence of 100 g/L sulphuric acid concentrated for treatment of Abu Rushdur sample to dissolve uranium in the presence of an oxidizing agent. The leaching efficiency of uranium reached to about 87% (Ahmed et al., 2015). Another experiment on Abu Rashid samples by M.S. Nagar et al. has shown that 91.5% and 52% uranium leachability for Caro's acid and dilute sulfuric acid respectively (Mohamed S. . Acid leaching of uranium from the El-Sela low-grade uranium deposit alters the ore particle size by dissolving the oxides that block joints and fractures, resulting in a drop in ore particle size and the swelling of small particles over time. Adding cobblestone to improve permeability would result in a 79% increase in leaching efficiency . The application of agitation and column percolation leaching techniques to uranium-rich mineralization (El Missikat) revealed that these procedures were successful in producing significant results. After 40 days of operation, 75% of leaching efficiency utilizing 24 kg/ton acid consumption, it was found that El Missikat uranium mineralized ore is practicable for leaching (M. S. Nagar et al., 2020). Using column experiments, the impact of various parameters on uranium leaching and acid consumption from Qash Amir uranium deposit was examined. At a flow rate of 10 L/m 2 /h, 74.2% uranium recovery was achieved after 40 days of column leaching trials, with 26.2 kg of acid consumed per ton of ore (Mohamed Soliman . Acid leaching of uranium from uraniferous granite sample assaying 150 mg/L of G-II has been studied for the recovery of uranium by the agitation leaching technique. Under the optimum conditions, it was possible to realize a dissolution efficiency of about 94% for uranium (Kraiz et al., 2016). Uranium was recovered by acid leaching from the waste tailing of uranium mineralization in (G-II) after re-crashing the uranium mineralized ore to-10 mm, using 20 g/L sulphuric acid solution. The leachability of uranium in the column under acid leaching conditions reached a maximum value of 83.9 (Gazala & Nagar, 2017). The small column also investigated after applying optimum conditions in G-II uranium mineralization, and the obtained results show that the leaching efficiency attained about 78.3% using 34 kg/ton at 44 days leaching period (MS Nagar et al., 2021).
The Yotsugi deposit of the Ningyo-Toge mine was explored for low grade and coarse uranium ore. It was discovered that leaching coarse and low-grade ore (−150 mm and 0.07% U) in a vat with sulphuric acid solutions yields a high uranium recovery, approximately 88%. The overall time required for leaching and washing was around 2 weeks. The height of the ore bed in the vat was used to calculate leach solution flow rates. The height of the ore bed had a significant impact on the flow rate of the leach solution (Sugitsue & Onda, 1981).
The aim of this paper is to investigate the sprinkling intensity and acid consumed during column and vat leaching of uranium mineralization from Gattar-II, South Eastern Desert, Egypt, as technoeconomic study.

Column leaching tests
The percolation leaching via column leaching tests was done using a representative sample from the uranium mineralization in Gattar II area. The leaching experiments were performed using four poly vinyl chloride (PVC) columns 100 mm in diameter and 1.5 m in height. A total of 15.0 kg of uranium mineralization sample was loaded into each column. The effect of the physical and chemical parameters was assessed and optimized using preliminary metallurgical experiments as column leaching using the leaching conditions shown Table 1. Using three particle size fractions of −40 mm, −20 mm, and −10 mm, the effect of particle size on the efficiency of uranium leaching was investigated as −10 mm (Mahmoud et al., 2001). For this study, an ore sample with a size of −10 mm was chosen.

Vat leaching experiment
The vat (box) leaching having a general slope of about 1% from front to back to allow solution collection. Its dimension, (5 m width *12 m length *5 height) the total ore amount about 350 ton. To prevent seepage of pregnant liquor, the ground surface of the pad covered with polyethylene sheets, and about 10 cm sand filter layer for the further filtration process. Driblet irrigation pipe is adopted to distribute lixiviant over the ore surface, to minimize the rate of evaporation. Mixing of concentrated H 2 SO 4 and water or barren solution (lixiviant), and the flow rate of a lixiviant into the heap is adjusted.

Chemical analyses of ore samples
Wet chemical procedures were used to evaluate the crushed samples (Shapiro & BRANNOCK, 1962). The concentrations of Na and K were determined using a flame photometric technique, whereas SiO 2 , Al 2 O 3 , TiO 2 , and P 2 O 5 were determined using a spectrophotometric method. Titration methods were used to determine total iron as Fe 2 O 3 , MgO, and CaO. The loss on ignition (L.O.I.) was calculated using gravimetric methods. The estimated inaccuracy for the key constituents was less than 1%. All the chemical analyses were carried out in the laboratories of the Nuclear Materials Authority, Egypt.

Uranium determination
The uranium contents of the ore-samples, intermediate pregnant solution, and pregnant leaching solutions were determined according to the method described by Davies and Gray and modified by Nuclear Materials Authority laboratories, Egypt (Davis and Gray, 1964).

Chemical characteristics of G-II mineralized sample
As a matter of the fact, the nature and content of major elements are generally quite important in selecting the leaching technique. Presence of acid consuming minerals makes the application of an alkaline method for uranium leaching more economic and vice versa. Acid leaching is generally preferred because it gives higher result for uranium dissolution, unless the ore contains minerals that would reduce the acid consumption prohibitively excessive. The chemical composition of G-II mineralized sample was found sufficiently simple and well situated for acid leaching since its assay is low in CaO (1.7%) and MgO (0.5%). In addition, the silicic gangue minerals (feldspars and quartz) equivalent to 73.3% SiO 2 are frequently predominating and which are practically inert to acid attack. The other major oxides, namely, Al 2 O 3 , Na 2 O, and K 2 O assaying 12.3, 3.5, and 3.4%, respectively, are mainly present in feldspar minerals except a relatively small Al 2 O 3 content in the clay minerals .The latter are indeed of minor importance as indicated from the low loss on ignition (1.6%).

Results of columns leaching of G-II uranium mineralization
To investigate the acid leaching of uranium mineralization in the course of heap leaching and evaluate the reagent consumption per unit mass of uranium dissolution. In this paper, the sprinkling intensity and acid consumed during column leaching are investigated. In the leaching experiments, lixivant was added until the dissolved uranium value decreased to a sufficiently almost steady low level. For this purpose, four column percolation leaching experiments have been performed under the conditions given in Table 1.

Acid consumption
Acid leaching refers to use sulfuric acid as lixiviant in all commercial practice. Sulfuric acid ionizes in solution to form sulfate, bisulfate, and hydrogen ions. Reaction with hexavalent uranium, which dissolves as UO 2 2+ cation, produces uranyl sulfate and complex uranyl sulfate anions as follows: The most important factor in the cost of chemical reagent treatment of uranium mineralization has been the amount of consumed sulfuric acid which produces a satisfactory recovery in the leaching stage. The characteristics of the feed uranium mineralization and the leaching conditions influence the acid concentration used in leaching. Excessive acid can speed up recovery that lead to increased acid consumption. Furthermore, the high acid concentration can cause an excessive amount of associated gangue minerals to dissolve in the pregnant solution. To avoid the precipitation of dissolved uranium, the (pH) in the leaching liquor should be maintained at a certain value.
As shown in Figure 1, the column with the acid concentration (40 g/L) had a higher uranium dissolution with a maximum recovery of 77.6% compared to 74.5% recovery in column with the acid concentration (20 g/L) for the same 42-day leaching period. Extending the leaching period in column (20 g/L) to 56 days increased total uranium recovery by 4.8% and increased the recovery rate to 79.3%.

Interpretation of acid consumption from techno-economic view
The required reagent consumption per percentage increase in uranium recovery within the ore sample for different acid concentrations is known as the leaching index (ẞ) (Ghorbani, Marcelo et al., 2016). As seen in the following equation: where (β) is the leaching index, V is the total volume of the consumed liquid (L), M is the total mas of the consumed acid (g), and E is the uranium dissolution (%).
The leaching index (β), as shown in Table 2 and Figure 1, was calculated for column (20 g/L) during the extension of leaching time from day 42 to day 56. According to the leaching index (β), 87.5 g acid is required to achieve a 1% increase in uranium recovery in column (20 g/L) after day 42, while the maximum amount of acid required up to day 42 per unit of recovery was around 18 g. After day 42, the leaching index for column (40 g/L) indicates that 400 g acid was consumed (if extended). While, as shown in Table 3, results show that 21.2 and 32 g of acid are consumed over the duration of the dissolution process, with acid concentrations of 20 and 40 g/L, respectively, to achieve a 1% increase in uranium recovery. 0.05:1, respectively. The time period for process and solution consumption are two key technoeconomic factors and play a key role in the process design and decision-making steps. From this standpoint, optimum sprinkling intensity 0.1:1 is preferable to 0.2:1.0 sprinkling intensity, the leaching rate is almost the same, but, when using 02.1 sprinkling intensity it will introduce solution reagent in the rate higher than in 0.1:1 and hence will decrease the PLS residence time, concentration, and correspondingly more of the lixiviant will be percolated through the column with little or no interaction with ore particles. As in Figure 2, extending the leaching time in column 2 (0.1:1) up to 56 days increased the overall uranium recovery by about 4% and the recovery reached 79.3%, while the lowest uranium recovery was obtained at sprinkling intensity of 0.05:1 during 62 days. The lixiviant's percolation rate should be slow enough to allow the lixiviant to have enough contact time with the ore particles to achieve a high rate of ore dissolution.

Accumulative liquid/solid ratio
In Figure 2, more than 50% of uranium dissolution during the first week of leaching in all sprinkling intensities, this is believed to be due to all fast-dissolution of uranium near the surface, and after that, the leach reagent has to migrate from the particle surface into the particle to recover further uranium. These results are consistent with what Ogbonna found in 2006 (Ogbonna et al., 2006). at different sprinkling ratios.
During the leaching experiment, a higher irrigation flow rate dilutes the uranium concentration in the pregnant solution of the column, as shown in  fresh solution and recycling into the leaching process has an positive impact on the uranium extraction rate at various sprinkling intensities in all time periods. The addition of fresh solution is more successful during the time periods of (14-21), (24-30), and (37-41) days, as shown in Figure 3.

Interpretation of reagent consumption from techno-economic view
As given in Table 4, the leaching index (β) has been calculated for column 2 during the extension of leaching time from day 47, up to day 56. The results show that after day 47, 3.1, and 23 L solution was needed to achieve a 1% increase in uranium recovery in column 2 and 3 with a sprinkling strength of 0.1:1 and 0.2:1, respectively, Figure 4 shows the average values of the leaching index (ẞ) vs. leaching time for different sprinkling intensities in column 2 (0.1:1) and column 3 (0.2:1). There is no linear effect of having a high irrigation rate in the column leaching. In fact, the process of dissolving uranium from the two columns has several stages. After the surface has been depleted of quick uranium dissolution, leaching would most likely continue to dissolve uranium mineralization values for slower leaching minerals near the surface, rather than the faster-leaching minerals closer to the center of large ore particles (Ghorbani et al., 2011).

Applying optimum conditions obtained from column leaching onto vat leaching process
Gattar uranium ore assaying 0.035%U was subjected to crushing to <50 mm grain size, then vat leached with diluted sulfuric acid (50~20 g/L) at optimized conditions, a leaching efficiency of about 70.7% was obtained. The obtained pregnant solution assayed 570 mg/L contains complex of

Interpretation of reagent consumption during vat leaching as a techno-economic view
3.3.1.1. Lixivient consumed. From the obtained data of vat leaching that extended for 90 days, it was found that a total volume of sprinkling lixivant attaining about 290 m 3 , in the meantime, an overall leaching efficiency of about 70.7% has been obtained and the accumulative liquid to solid ratio attain about 0.83, as shown in Figure 5.

Recycling solutions.
The obtained leach liquor was recycled in the same vat (350 ton) in order to obtain relatively enriched pregnant liquor. In Figure 6, the recycling period at (1-10), (28-32), (48-52), and (63-66), the uranium in pregnant solution increased from (250-600), (360-500), (150-350), and (120-210) ppm, respectively. Another reason for using the recycling process is to adjust the pH of pregnant uranium solution (in the first 10 days of leaching). The most important reason for the recycling process is to save reagent consumption, as one-third of the consumed solutions is saved during the recycling process.

Lixiviant (barren) regeneration.
Leach solution management is an important part of heap leach operation, both to maximize metal recovery and to prevent the loss of reagent bearing solution to the environment. The resulting acidic barren solution from ion exchange (IX) process is recycled back into the vat via the barren pond. Taking into account that, the first 10 bed volumes of the adsorption cycle are wasted to tailing in order to minimize chloride ions that obtained from elution process . The resulting PLS often includes different types of the desired metal, especially  the types of decomposed impurities that were not recovered in the IX stage. The increased element concentrations increase the density, impurities and viscosity of the process solutions, and form insoluble secondary products that may be precipitated through the solution channels and/or coated resin surfaces.
The resulted effluent solution from the adsorption process in G-II was analyzed for determining the associated elements which are, Mn, Zn, Na, Mg, Fe, and Si. These ions are less adsorbed on resin during uranium adsorption since their effluent /leach percent being 95.3,76,81.1,76.5, 90, and 74.6%, respectively. The barren solution was recycled to leaching for a total of 4 cycles. A laboratory test verified that uranium was dissolved in the recycled solution in spite of the sulfuric acid concentration (15 g/L). In addition, the solution contains a certain concentration of ferric ion (1.9 g/L) which speeds up the process of dissolving uranium dissolution.
From the leaching index (β), the result indicates that 4.1 m 3 solutions are consumed to have a 1% increase in the uranium recovery of vat leaching with the sprinkling intensity of 0.1:1 after 90 day, as shown in Table 5.
3.3.1.4. Acid consumption during vat leaching. Acid consumption is the major economic factor in uranium extraction from uranium ores by vat leaching technique. As shown in Figure 7, the increase in the uranium dissolution with a maximum recovery of 70.7% was obtained after the leaching period of 90 days. At the same time, the overall consumed acid after this period was 17.6 kg/ton of ore.
From the leaching index (β) in Table 6, the result indicates that 88 kg acid is consumed to have a 1% increase in the uranium recovery of vat leaching with the acid concentration between 50-10 g/L after 90 days. Finally, a proposed flow sheet for the leaching of uranium mineralization is shown in Figure 8.

Conclusion
Circulation of the intermediate leaching solutions (ILS) and barren solutions have a notable effect on the capital cost of the heap leaching system. The most important reason for the recycling process is to save reagent consumption, as one-third of the consumed solutions are saved during the recycling process. Results indicate that in column leaching 21.2 and 32 g of acid are consumed using an acid concentration of 20 and 40 g/L, respectively, to have a 1% increase in the uranium recovery. Also, 1.0 and 1.8 L solutions using 0.1:1 and 0.2:1 sprinkling intensities are consumed respectively to have a 1% increase in the uranium recovery. The results of vat leaching indicate that 4.1 m 3 solution and 88 kg acid are consumed to have a 1% increase in the uranium recovery of vat leaching with the sprinkling intensity of 0.1:1 after 90 days. At the same time, the overall consumed acid after this period was 17.6 kg/ton of ore. In the future, these parameters can be used to process the uranium mineralization of the Gattar area through the heap leaching at the lowest possible economic cost.