Figure
1.
Flow chart of comprehensive utilization of zinc oxidized ore.
Xiao-yi Shen, Hong-mei Shao, Ji-wen Ding, Yan Liu, Hui-min Gu, and Yu-chun Zhai, Zinc extraction from zinc oxidized ore using (NH4)2SO4 roasting−leaching process, Int. J. Miner. Metall. Mater., 27(2020), No. 11, pp.1471-1481. https://dx.doi.org/10.1007/s12613-020-2015-2
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An improved method of (NH4)2SO4 roasting followed by water leaching to utilize zinc oxidized ores was studied. The operating parameters were obtained by investigating the effects of the molar ratio of (NH4)2SO4 to zinc, roasting temperature, and holding time on zinc extraction. The roasting process followed the chemical reaction control mechanism with the apparent activation energy value of 41.74 kJ·mol−1. The transformation of mineral phases in roasting was identified by X-ray diffraction analysis combined with thermogravimetry–differential thermal analysis curves. The water leaching conditions, including the leaching temperature, leaching time, stirring velocity, and liquid-to-solid ratio, were discussed, and the leaching kinetics was studied. The reaction rate was obtained under outer diffusion without product layer control; the values of the apparent activation energy for two stages were 4.12 and 8.19 kJ·mol−1. The maximum zinc extraction ratio reached 96% while the efficiency of iron extraction was approximately 32% under appropriate conditions. This work offers an effective method for the comprehensive use of zinc oxidized ores.
Zinc is an important nonferrous metal that is extensively applied to galvanization, alloys, batteries, and other fields. Sulfide ores have long been the main raw material in zinc metallurgy [1−2]. However, due to the overexploitation of sulfide ores and the increased zinc demand, much attention has been paid to the reasonable exploitation of zinc oxidized ores [2−9].
As the largest resources bearing zinc [4−5,10], zinc oxidized ores usually exist as oxidized carbonate or silicate minerals, such as smithsonite (ZnCO3), willemite (ZnSiO4), and hemimorphite (Zn4Si2O7(OH)2·H2O) [4,11−14]; in addition, they usually contain high-grade silica [7,15]. Extensive studies have been carried out to explore the treatment of zinc oxidized ores. Flotation has been applied to zinc oxidized ores, but its use is difficult because of the fine intergrowth, complex phase compositions, and high gangue content of zinc oxidized ores [3,16−17]. Pyrometallurgical and hydrometallurgical routes are adopted to utilize zinc oxidized ores. Traditional pyrometallurgical processes are not highly competitive because of their high energy consumption and high CO2 and residue emissions [17−21]. Hydrometallurgical routes are generally divided into acid leaching and alkaline leaching, which includes ammonia leaching and sodium hydroxide leaching. Sulfuric acid leaching is widely applied and remarkably versatile [19,22]. However, the production parameters in this process require proper control to prevent the formation of silica gel, which hampers filtration [6−7,23−24]. Several attempts have been made to improve filtration performance, and they include precipitating silica, controlling pH values [25−26], adding flocculating agents [12,26], and adding Al3+ [11,24,27]. Ammonia leaching has high efficiency in treating zinc oxidized ores for the formation of stable zinc ammine complexes [2,9,12,26]. In this process, leaching vessels need to be hermetically sealed to avoid ammonia volatilization [3,17,28]. Sodium hydroxide leaching is also a promising route, and many published reports have referred to the alkaline treatment of zinc-bearing minerals. However, the decomposition of willemite and hemimorphite is slow [7,29−31], and additional work is needed to realize the separation of Zn, Si, and Pb [7,32]. Bioleaching has also been carried out, but it is far from realizing industrialization [33]. As a strong acid and weak base salt, ammonium sulfate is frequently applied in metallurgy for the extraction of valuable metals from various low-grade ores, including zinc oxidized ores, wollastonite, and blast furnace slag [10,34−37]. An improved and effective process was proposed to realize the comprehensive utilization of zinc oxidized ores. First, a mixture of zinc oxidized ore and (NH4)2SO4 was roasted. After water leaching and filtration, zinc and small amounts of iron entered the filtrate and were separated from the silica and calcium that remained in the leaching residue. The roasting gas absorbed by dilute H2SO4 or water was transformed into (NH4)2SO4 or ammonia water [17]. Second, the leaching solution was used to prepare zinc products after purification, and iron was utilized subsequently. Third, the enriched strontium and lead in the leaching residue were recovered through a conversion method. Fourth, silica was extracted and used to prepare white carbon black, fibrous xonotlite, and stereo-porous tobermorite. Finally, the remaining residue was used to recover iron. In the whole process, valuable components, including Zn, Si, Pb, Sr, and Fe were extracted and made into products. Water, (NH4)2SO4, and NaOH were cycled. The flow chart of the proposed process is shown in Fig. 1.
In this investigation focusing on zinc extraction, a typical zinc oxidized ore from Lanping located in southwest China was used as a raw material. It was roasted by industrial (NH4)2SO4 and then subjected to water leaching. The effects of the molar ratio of (NH4)2SO4 to zinc, roasting temperature, and holding time on the extractions of zinc and iron were determined. The transformation of mineral phases was identified by X-ray diffraction (XRD) and thermogravimetry–differential thermal analysis (TG–DTA). Moreover, the effects of water leaching temperature, time, stirring velocity, and liquid-to-solid (L/S) ratio on the extraction ratio of zinc were discussed. The roasting mechanism and water leaching kinetics were determined using the constant conversion method and shrinking core model, respectively.
The zinc oxidized ore with a particle size of less than 74 μm was used as raw material. Industrial grade (NH4)2SO4 serving as the reactant was used as received.
A resistance furnace (temperature accuracy of ±1°C) was employed in the roasting experiments. Exactly 50 g of powder-like zinc oxidized ore and a specific amount of (NH4)2SO4 were uniformly mixed and put in a crucible that was then placed in the furnace. After the temperature reached the specified value ranging from 300 to 500°C and maintained within 2.5 h, the sample was taken out and leached in water at 80°C for 1 h with an L/S ratio of 4:1 mL/g. Once the temperature reached the required value, the mixtures were removed within a predesigned time interval of 10 min and then cooled rapidly for the kinetic analysis.
The leaching investigation was conducted at a certain L/S ratio ranging from 2:1 to 6:1 mL/g. A vessel containing distilled water in a water bath was continuously stirred. Once the temperature rose to the desired value of 20 to 95°C and stabilized, the samples were added into the vessel and leached for a period of time under stirring. The slurry was then filtrated before analysis. For the kinetic analysis, the specimens were collected at 5 min intervals, filtered, and determined using ethylenediaminetetraacetic acid titration.
The extraction ratio of zinc was determined using the following equation:
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(1) |
where η is the extraction ratio of zinc; V is the volume of the leaching solution, L; c is the zinc concentration, g·L−1; G is the weight of the zinc oxidized ore, g; and w is the level of zinc in the zinc oxidized ore.
The main compositions of the zinc oxidized ore analyzed by the chemical method are listed in Table 1. The powder-like ore was characterized by XRD and scanning electron microscopy (SEM), as shown in Fig. 2. The content of ZnO, as one of the main components, was 29.12wt%. It mainly existed in the form of ZnCO3 and Zn2SiO4. The contents of PbO and SrO were 3.66wt% and 1.81wt%, respectively. The other components were 9.02wt% iron oxide, 27.85wt% SiO2, and 2.50% CaO. The XRD study indicated that smithsonite (ZnCO3) and quartz (SiO2) existed as the major mineral phases and that CaCO3, PbCO3, CaSO4·2H2O, Fe2O3, and Zn2SiO4 existed as the minor phases. The zinc oxidized ore particles were uneven and irregular.
| ZnO | Fe2O3 | SiO2 | PbO | SrO | CaO | CO2 | SO3 | Al2O3 | Others |
| 29.12 | 9.02 | 27.85 | 3.66 | 1.81 | 2.50 | 15.38 | 2.13 | 3.22 | 5.31 |
Experiments were performed to examine the effect of the molar ratio of (NH4)2SO4 to zinc, roasting temperature, and holding time on the extractions of zinc and iron at particle sizes of less than 74 μm. The results are plotted in Fig. 3.
This study investigated the effect of the molar ratios of (NH4)2SO4 to zinc in the zinc oxidized ore in the range of 1.1:1–1.45:1 on the extractions of zinc and iron at 450°C, which was maintained for 2 h (Fig. 3(a)). An increase in the molar ratio exerted an appreciable impact on the extractions of zinc and iron. The zinc extraction ratio increased from 74.38% at a molar ratio of 1.1:1 to 94.99% at a molar ratio of 1.4:1. The extraction ratio is close to that of pressure acid leaching reported by He et al. [24]. Thereafter, the zinc extraction ratio showed no obvious increase, but the iron extraction increased continuously. Sufficient (NH4)2SO4 is necessary for obtaining a high zinc extraction ratio because it improves the interaction between reactants by enhancing the contact area as the dosage of (NH4)2SO4 increases. However, excessive (NH4)2SO4 results in not only the excess consumption of auxiliary material but also a high reaction ratio of iron. The molar ratio of 1.4:1 was chosen in this study.
The study on the effect of roasting temperature ranging from 300 to 500°C on the extractions of zinc and iron was performed under a molar ratio of (NH4)2SO4 to zinc of 1.4:1 and holding time of 2 h. The data shown in Fig. 3(b) reveal that the roasting temperature exerted an important effect on the roasting process. The zinc extraction ratio increased from 10.09% at 300°C to 94.99% at 450°C. However, the variation was minimal at temperatures above 450°C. This result implies that the temperature of 450°C is appropriate for extracting zinc. Furthermore, the iron extraction ratio continuously increased. The increasing temperature promoted the reaction. This result is explained as follows: (NH4)2SO4 is first decomposed into NH4HSO4 and NH3 and then into (NH4)2S2O7, and even into N2, SO2, and NH3 [38−42]. The reaction is transformed from the solid–solid phase to the liquid–solid phase under the roasting temperature, which presumably decreases the reaction resistance [38−42].
The effect of holding time on the extractions of zinc and iron was investigated at a roasting temperature of 450°C and molar ratio of (NH4)2SO4 to zinc of 1.4:1. The zinc extraction ratio shown in Fig. 3(c) increased obviously within 1 h; thereafter, its change was inappreciable. Zinc extraction ratio was 96.11% under roasting for 1 h while the iron extraction ratio was 32.49%. This result implies that the holding time of 1 h is enough for extracting zinc. Iron extraction ratio continuously increased. Prolonging holding time increases the burden of eliminating iron and decreases the extraction efficiency. To reduce energy consumption, this study adopted the holding time of 1 h.
The appropriate roasting conditions could then be concluded as follows: roasting temperature of 450°C, molar ratio of (NH4)2SO4 to zinc in zinc oxidized ore of 1.4:1, and holding time of 1 h at particle sizes less than 74 μm.
Fig. 4 shows the representative TG−DTA curves ranging from 25 to 600°C of (NH4)2SO4 and the mixture of zinc oxidized ore and (NH4)2SO4 obtained at 10°C·min−1 of heating rate and 100 mL·min−1 of air flow rate. Three obvious endothermic peaks appeared at approximately 310, 336, and 435°C (Fig. 4(a)). These peaks were accompanied by three prominent weight losses in the TG curve of 12.80wt%, 6.01wt%, and 80.77wt% at the temperature ranges of 250−325°C, 325–350°C, and 350–440°C, respectively. The weight losses of 12.80% and 6.01% were extremely close to the stoichiometric weight loss of losing NH3 from (NH4)2SO4 and the formation of (NH4)2S2O7. The three endothermic peaks were ascribed to the decomposition of (NH4)2SO4, NH4HSO4, and (NH4)2S2O7 according to the following respective reactions [38−39]:
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(2) |
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(3) |
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(4) |
Five obvious endothermic peaks appeared at approximately 60, 205, 302, 335, and 420°C (Fig. 4(b)). These peaks were accompanied by three prominent weight losses in the TG curve of 11.16%, 12.81%, and 21.51% at temperature ranges of room temperature–280°C, 280–320°C, and 320–440°C, respectively. The first stage of weight loss was due to the dehydration and release of CO2 from ZnCO3. The second stage was mainly attributed to (NH4)2SO4 decomposition. The third stage was extremely complex as it involved the formations followed by the deamination of (NH4)2Zn(SO4)2 and (NH4)3Fe(SO4)3, the formations of CaSO4 and PbSO4, and the total decomposition of (NH4)2SO4 [38−40].
To shed light on the roasting process, this work used XRD testing in identifying the phase structures of the specimens obtained at different roasting temperatures (Fig. 5). The main phases in the specimen obtained at 250°C were almost the same as those of the initial mixture. This result indicated that no substantial reaction occurred due to the absence of strong diffraction peaks of (NH4)2SO4, Zn2SiO4, and PbCO3. The main phases in the specimen roasted at 300°C were ZnCO3, SiO2, (NH4)2SO4, Zn2SiO4, (NH4)2Zn(SO4)2, and NH4HSO4. (NH4)2Zn(SO4)2 was synthesized, and the decomposition or reaction of (NH4)2SO4 was incomplete. In the specimen obtained at 350°C, ZnCO3 and (NH4)2SO4 disappeared, and CaCO3 remained stable. This result indicated the completion of the decomposition or reaction of ZnCO3. (NH4)3Fe(SO4)3 formed at 350°C (Fig. 5(d)).
The main phases in the specimen obtained at 400°C were CaSO4, SiO2, (NH4)2Zn(SO4)2, PbSO4, CaCO3, NH4Fe(SO4)3, and silicate. Hence, CaSO4 and PbSO4 were formed, and NH4Fe(SO4)2 was generated from (NH4)3Fe(SO4)3. The main phases in the specimen obtained at 425°C were SiO2, ZnSO4, CaSO4, (NH4)2Zn(SO4)2, PbSO4, NH4Fe(SO4)3, and Fe2(SO4)3. ZnSO4 was ascribed to the decomposition of (NH4)2Zn(SO4)2. CaCO3 disappeared, thus revealing the complete transformation to CaSO4. Fe2(SO4)3 was generated from NH4Fe(SO4)2. The main phases in the sample obtained at 450°C were similar to those of the specimen obtained at 425°C. In addition, (NH4)2Zn(SO4)2 disappeared, and the intensity of the diffraction peaks of ZnSO4 and PbSO4 increased. Fe2(SO4)3·9H2O was attributed to the moisture absorption of the roasted specimen. Quartz SiO2 was stable throughout the whole roasting process. The diffraction peaks of NH4HSO4 are not conspicuous in Figs. 5(c) and 5(d), but NH4HSO4 was clearly identified as the decomposition product of (NH4)2SO4 in some reports [34−37].
The phase transformation in the roasting process can be summarized as follows: ZnCO3 and its decomposition derivative ZnO transform into (NH4)2Zn(SO4)2 at 300°C and further transforms into ZnSO4 at 425°C; Fe2O3 transforms into (NH4)3Fe(SO4)3 at 350°C and further changes into NH4Fe(SO4)2 at 400°C and Fe2(SO4)3 at 425–450°C; PbSO4 and CaSO4 are obtained when the temperature reaches 400°C.
The XRD patterns and TG−DTA curves revealed that the peak of the TG−DTA curves at 302°C was mainly caused by the decomposition of (NH4)2SO4 and the formation of (NH4)2Zn(SO4)2. Sun et al. [17] proved that (NH4)2SO4 begins to decompose into NH4HSO4 when the temperature exceeds 260°C. In the current work, the peak at 335°C was mainly ascribed to the decomposition of NH4HSO4 and the formation of (NH4)3Fe(SO4)3. The peak at 420°C was mainly attributed to the simultaneous and continuous formations, followed by the deamination of (NH4)2Zn(SO4)2 and (NH4)3Fe(SO4)3 [40] and by the formations of CaSO4 and PbSO4. Li et al. [39] studied the decomposition process of (NH4)2SO4 in detail and presented the TG−DTA curves and Fourier-transform infrared (FT-IR) spectra. The results showed that a large amount of NH3 was released at 384°C and that a large amount of SO2 and H2O was produced at 520°C. Zhang et al. [38] presented the coincident TG−DTA curves obtained in argon protection, and Yin et al. [41] reported the homologous FT-IR spectra; the most significant weight loss occurred at the temperature range of 450–520°C due to the total decomposition of (NH4)2SO4 (Eq. (4)). Yin et al. [41] also pointed out that the decomposition of ammonium bisulfate is more difficult than that of (NH4)2SO4 and that the reaction rate of ammonium bisulfate with blast furnace slag is extremely fast. The chemical reactions that mainly occurred before the total decomposition of (NH4)2SO4 may be deduced as follows:
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(5) |
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(6) |
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(7) |
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(8) |
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(9) |
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(10) |
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(11) |
As discussed above, (NH4)2SO4 was decomposed into NH3 and NH4HSO4 before the total decomposition occurred at temperature range of 350–450°C. The reaction was a gas–liquid–solid reaction; thus, the shrinking unreacted core model was not suitable for kinetic analysis. The constant conversion method was thus adopted [41]:
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(12) |
where t is the reaction time, min; A is the frequency factor, min−1; Ea is the apparent activation energy, kJ·mol−1; R is the molar gas constant, J∙mol−1·K−1; and T is the thermodynamic temperature, K.
The results in Fig. 6(a) show that with the increase in roasting temperature and time, the zinc extraction ratio increased gradually. The roasting mechanism was studied, and the plot of ln(1/t) versus 1/T is shown in Fig. 6(b). The calculated value of the apparent activation energy from the slopes of the fitting lines was 41.74 kJ·mol−1, which falls in the typical range of chemical reaction control mechanism [43].
The effects of leaching temperature, time, stirring velocity, and L/S ratio on the extraction ratio of zinc were investigated, and the results are plotted in Fig. 7.
The experiments to determine the effect of leaching temperature ranging from 20 to 95°C were performed under the following conditions: L/S ratio of 4:1 mL/g, leaching time of 60 min, and stirring velocity of 400 r·min−1. The extraction ratio of zinc increased with temperature (Fig. 7(a)). Increasing temperature enhanced the dissolution of ZnSO4 due to the improvement of molecular movement. After 90°C, the zinc dissolution began to decline slightly. This result was attributed to the synthesis of ammonium jarosite, entraining Zn2+ as an inclusion [44].
The study on the effect of leaching time was performed under the following conditions: L/S ratio of 4:1 mL/g, leaching temperature of 90°C, and stirring velocity of 400 r·min−1. Prolonging leaching time increased the extraction ratio of zinc within 60 min (Fig. 7(b)). However, the increase was minimal after 50 min. The following experiments were carried out for 50 min.
The stirring velocity experiments ranging from 100 to 600 r·min−1 were performed under the following conditions: L/S ratio of 4:1 mL/g, leaching temperature of 90°C, and leaching time of 50 min. The extraction ratio of zinc increased obviously with the stirring velocity ranging from 100 to 400 r·min−1 (Fig. 7(c)) due to the improvement of the relative motion of the liquid–solid and the enhancement of mass transfer. However, the change was minimal when the stirring velocity varied from 400 to 600 r·min−1. This result indicated that the effect of stirring velocity was relatively weak. Thus, 400 r·min−1 was selected in the following experiments.
The impact of L/S ratio on extraction ratio of zinc was studied at 90°C and stirring for 50 min at 400 r·min−1. The results plotted in Fig. 7(d) indicate that the extraction ratio of zinc increased with the L/S ratio rising within 4:1 mL/g because of the decrease of the ion concentration in water. Thereafter, the extraction ratio of zinc tended to be stable. Therefore, the L/S ratio of 4:1 mL/g was chosen.
The extraction ratio of zinc against time (0–60 min) was examined at different temperatures within 313–353 K (40–80°C), and the results are plotted in Fig. 8(a). The extraction ratio of zinc increased steadily with temperature and time. Approximately 93% zinc was leached at 353 K over 40 min. The water extraction process was a typical liquid–solid reaction without a solid product layer formation. Thus, the following shrinking core model (Eq. (13)) was used to explain the kinetic mechanism.
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(13) |
where α is the reaction fraction; k is the diffusion rate constant, min–1; and t is the reaction time, min.
The leaching process can be divided into two stages of 0–5 min and 5–30 min. The kinetics data in Fig. 8(b) (0–5 min) and Fig. 8(c) (5–30 min) agreed well with Eq. (13). The Arrhenius plots of lnk versus 1/T are presented in Figs. 8(d) and 8(e). The activation energy and pre-exponential factor was calculated, and the values were 4.12 kJ·mol−1 and 0.3490 min−1 for the first stage and 8.19 kJ·mol−1 and 0.0963 min−1 for the second stage, respectively. Thus, the reaction rate equations of the two stages (0–5 min and 5–30 min) conforming outer diffusion mechanism without product layer are derived as follows:
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(14) |
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(15) |
The main components of the residue analyzed by the chemical method are listed in Table 2. The contents of ZnO, Fe2O3, and SiO2 were 1.94wt%, 12.02wt%, and 51.82wt%, respectively. PbO and SrO were enriched to 6.81wt% and 3.38wt%, respectively. The leaching residue is a valuable resource bearing Pb and Sr and is worth utilizing. The XRD pattern and SEM image presented in Fig. 9 indicate that the main phases in the residue were SiO2, PbSO4, SrSO4, Fe3O4, and CaSO4. The residue particles were irregular and showed a rough surface.
| ZnO | Fe2O3 | SiO2 | PbO | SrO | CaO | SO3 | Others |
| 1.94 | 12.02 | 51.82 | 6.81 | 3.38 | 4.67 | 14.32 | 5.04 |
(1) The effects of roasting and leaching parameters on zinc extraction were evaluated. The extraction ratio of zinc reached the maximum of 96% under the following operating conditions: molar ratio of (NH4)2SO4 to zinc of 1.4:1, roasting temperature of 450°C, holding time of 1 h, leaching temperature of 90°C, time of 50 min, stirring velocity of 400 r·min−1, and L/S ratio of 4:1 mL/g.
(2) The kinetic study demonstrated that the roasting process followed the chemical reaction control mechanism with the apparent activation energy value of 41.74 kJ·mol−1. The leaching control step involved the outer diffusion without product layer with the apparent activation energy values of 4.12 and 8.19 kJ·mol−1 of two stages.
(3) Ammonium sulfate roasting followed by water leaching can be adopted to extract zinc from zinc oxidized ores and is effective in practice. (NH4)2SO4 can be cyclically utilized in the whole process. The process may be used an alternative in dealing with zinc oxidized ores.
This work was financially supported by the National Natural Science Foundation of China (Nos. 51774070, 52004165, and 51574084), and the National Key Research and Development Program of China (No. 2017YFB0305401).
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必应学术
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| ZnO | Fe2O3 | SiO2 | PbO | SrO | CaO | CO2 | SO3 | Al2O3 | Others |
| 29.12 | 9.02 | 27.85 | 3.66 | 1.81 | 2.50 | 15.38 | 2.13 | 3.22 | 5.31 |
| ZnO | Fe2O3 | SiO2 | PbO | SrO | CaO | SO3 | Others |
| 1.94 | 12.02 | 51.82 | 6.81 | 3.38 | 4.67 | 14.32 | 5.04 |
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