- Email: [email protected]
A novel method to recover zinc and iron from zinc leaching residue
Minerals Engineering 55 (2014) 103–110
Contents lists available at ScienceDirect
Minerals Engineering journal homepage: www.elsevier.com/locate/mineng
A novel method to recover zinc and iron from zinc leaching residue Huan Yan a, Li-yuan Chai a,b,⇑, Bing Peng a,b, Mi Li a, Ning Peng a, Dong-ke Hou a a b
Institute of Environmental Science and Engineering, School of Metallurgy and Environment, Central South University, 410083 Changsha, Hunan, China Chinese National Engineering Research Center for Control & Treatment of Heavy Metal Pollution, 410083 Changsha, Hunan, China
a r t i c l e
i n f o
Article history: Received 18 June 2013 Accepted 26 September 2013 Available online 24 October 2013 Keywords: Zinc leaching residue Zinc ferrite Reduction roasting Acid leaching Magnetic separation
a b s t r a c t A novel method to recover zinc and iron from zinc leaching residue (ZLR) by the combination of reduction roasting, acid leaching and magnetic separation was proposed. Zinc ferrite in the ZLR was selectively transformed to ZnO and Fe3O4 under CO, CO2 and Ar atmosphere. Subsequently, acid leaching was carried out to dissolve zinc from reduced ZLR while iron was left in the residue and recovered by magnetic separation. The mineralogical changes of ZLR during the processes were characterized by XRF, TG, XRD, SEM–EDS and VSM. The effects of roasting and leaching conditions were investigated with the optimum conditions obtained as follows: roasted at 750 °C for 90 min with 8% CO and CO/CO + CO2 ratio at 30%; leached at 35 °C for 60 min with 90 g/l sulfuric acid and liquid to solid ratio at 10:1. The iron was recovered by magnetic separation with magnetic intensity at 1160 G for 20 min. Under the optimum operation, 61.38% of zinc was recovered and 80.9% of iron recovery was achieved. This novel method not only realized the simultaneous recovery of zinc and iron but also solved the environmental problem caused by the storage of massive ZLR. Ó 2013 Elsevier Ltd. All rights reserved.
1. Introduction Currently, more than 85% of the world’s metal zinc is produced by the conventional hydrometallurgical approach including oxidative roasting, acid leaching, purification, and electrowinning processes (RLPE) (Jha et al., 2001; Turan et al., 2004). During the oxidative roasting process, ZnS is converted to ZnO, but a significant fraction reacts with the iron impurities to form zinc ferrite (Dimitrova et al., 2000; Langová et al., 2009; Peng et al., 2012; Vahidi et al., 2009). Zinc ferrite is hardly soluble in mild acidic conditions, and a huge quantity of leaching residue is produced in the subsequent leaching process. The residues could be used to recover metallic values or it may be disposed off (Soner Altundog˘an et al., 1998). Pyrometallurgical and hydrometallurgical processes have been studied to recycle zinc and iron from ZLR. The most typical pyrometallurgical method is Waelze process, and the recovery of zinc could reach to 90–95% in this process. However, Waelze process consumes huge energy to maintain the high reaction temperature (above 1300 K), and large amount of high iron-bearing secondary residue is generated (Besße et al., 2010). These shortcomings also exist in other pyrometallurgical methods including Ausmelt technology (Hoang et al., 2009), roasting with Na2CO3 (Holloway
⇑ Corresponding author at: Institute of Environmental Science and Engineering, School of Metallurgy and Environment, Central South University, 410083 Changsha, Hunan, China. E-mail address: [email protected] (L.-y. Chai). 0892-6875/$ - see front matter Ó 2013 Elsevier Ltd. All rights reserved. http://dx.doi.org/10.1016/j.mineng.2013.09.015
et al., 2007), roasting with Na2S (Zhang et al., 2011), and some other methods (Antrekowitsch, 2001; Besße et al., 2010; Çoruh and Nuri, 2010). The most common hydrometallurgical process is to recycle zinc and iron from ZLR in a hot concentrated sulfuric acid followed by precipitation of the dissolved iron values from solution as jarosite, goethite or hematite (Graydon, 1988). High leaching rate both of zinc and iron could be obtained in these processes, but the removal of iron and other unwanted metals from the leaching solution is difficult (Dimitrios Filippou, 1997). Hydrochloric acid leaching (Langová et al., 2009), alkaline solution leaching (Dutra et al., 2006; Zhao and Stanforth, 2000), microwave caustic leaching (Xia and Picklesi, 2000), D2EHPA leaching (Vahidi et al., 2009), and leaching with various solvents have been studied (Langová et al., 2007). These hydrometallurgical processes are more economical because of lower capital and operating costs, but the purification process becomes more complex. Therefore finding a cost-effective and environmental-friendly process to recover zinc and iron from zinc leaching residue remains the major challenge. In this study, a simultaneously recycling zinc and iron from ZLR by the combination of selective reduction, acid leaching and magnetic separation is investigated. The objective of this recycling technology is to decompose ZnFe2O4 to ZnO and Fe3O4 selectively. Then, ZnO and Fe3O4 could be separated by acid leaching and magnetic separation, respectively. This study focus on the various operating parameters in the three stages under laboratory conditions to control phase transformation and search the optimum experiments conditions. The phase transformation, morphology variation and magnetism changes of the samples are detected by
104
H. Yan et al. / Minerals Engineering 55 (2014) 103–110
XRD, SEM and VSM to interpret some phenomenon observed in experiments and the decomposition procedure of ZnFe2O4.
500
2. Experimental approach
400
2 1
1----ZnFe2O4
6----Zn2SiO4
2----Fe3O4
7----PbS 8----Pb4SiO6
2.1. Materials and analysis ZLR used in this study was obtained from a zinc hydrometallurgical plant in Hunan province, China. The samples were dried at 105 °C for 24 h before grinding and sieved to obtain required fractions. Elements composition of the residue was determined by XRF and results in Table 1 show that the main contents in ZLR used in this study are Zn-19.57%, Fe-23.91%, S-6.40%, Si-4.47%, Pb-4.35%, Ca-2.21%. The crystalline phases of the samples were investigated by X-ray powder diffraction (Rigaku, TTR-III) in 2h scale using Cu Ka radiation (k = 1.5406 Å, 50 kV and 100 mA) at the scanning rate of 10°/min vary from 10° to 80°. As Fig. 1 shows that zinc ferrite (ZnFe2O4), magnetite (Fe3O4), zinc sulfide (ZnS), zinc silicate (Zn2SiO4), and calcium sulfate (CaSO4) are the main crystal mineral phases in the residue. The morphological changes were detected by Scanning Electron Microscopy (SEM, JEOL.LTD, JSM-6360LV). Vibration sample magnetometer (VSM) (Model: HH-15) was applied to assess the magnetism of magnetic concentrates. Potassium dichromate titration and EDTA titration methods were used to quantitative analysis the proportion of ferrous in total iron and content of zinc oxide to evaluate the effect of experiments (Zhang, 1992). 2.2. Experimental procedure 2.2.1. Reduction roasting The experiments were performed according to the flow sheet as shown in Fig. 2. Reduction roasting process was firstly operated in thermo-gravimetric apparatus (STA449F3) to decompose zinc ferrite to ZnO and Fe3O4 selectively. The ZLR was flatted on the inner wall of the silica crucible and pick up the suction. Then, the sample was heat up to determined temperature at 10 K/min under 100 ml/min Ar. As soon as the temperature reached, a proportion fraction of reduction gases (CO + CO2 + Ar) was added to take place
Intensity (counts)
3----CaSO4
9----FeS
4----PbSO4 5----ZnS
300 200 7
2 1
3
100
2 1
0 10
6 5 9 4 4 74 5
20
2 1
2 1 4
30
40
2 1
2 1 2 65 8 1
2 2 1 1
50
60
70
80
Angle (2-Theta deg.) Fig. 1. X-ray diffraction pattern of ZLR.
Table 1 Chemical composition of ZLR by XRF. Elements
Content (wt.%)
Elements
Content (wt.%)
Element
Content (wt.%)
Fe Zn S Si Pb Ca Mn Al
23.91 19.57 6.40 4.47 4.35 2.21 1.53 1.15
Cu As Cd Mg K Ba Sr Sn
0.8 0.52 0.31 0.26 0.24 0.20 0.16 0.16
Cl Ti In P Ag Cr Ni Mo
0.067 0.057 0.054 0.043 0.037 0.018 0.011 0.010
of Ar at 100 ml/min, respectively. After reduction roasting, the mixed reduction gases (CO + CO2 + Ar) were changed to Ar again to cool down the samples. The reduced products were stored in a tightly closed jar waiting for analysis. The influences of reduction roasting parameters such as CO concentration, CO/CO + CO2 ratio, roasting temperature and duration time were studied as detailed shown in Table 2. The content of zinc
Fig. 2. Flow sheet of the process of simultaneously recovering zinc and iron from ZLR.
105
H. Yan et al. / Minerals Engineering 55 (2014) 103–110 Table 2 Conditions used in each operation. Operations
Parameters
a
b
c
d
Reduction roasting
CO (%) CO/CO + CO2 (%) T (°C) t (min)
2–10 100 750 120
8 15–50 750 120
8 20 660–810 120
8 20 750 30–180
Acid leaching
T (°C) Acidity (g/L) S/L t (min)
25–65 90 9:1 30
35 70–120 9:1 30
35 90 5:1–13:1 30
35 90 9:1 10–60
Magnetic separation
t (min) Current intensity (G)
20 1160
– –
– –
– –
oxide and Fe2+/TFe ratio in reduced products were used to evaluate the reduction roasting effects. 2.2.2. Acid leaching Sulfuric acid leaching was conducted to selectively separate ZnO from the reduced ZLR. The leaching process was carried out in an agitator with which the leaching temperature and the speed of agitation can be controlled. The range of acidity, leaching temperature, leaching time and liquid to solid ratio in each operation were listed in Table 2. The remove rate of zinc and iron was used to evaluate the separation rate of zinc and iron. 2.2.3. Magnetic separation Magnetic separation was performed on a slurry of leaching residue (10% in solid) using a Davis tube magnetic separator (XCGS). Firstly, a certain amount of samples were milled with ball material
ratio at 10:1 and rotational speed at 300 r/min for 15 min to disperse the particles. Then, the magnetic separation was carried out to recover iron with magnetic intensity at 1160 G. After magnetic separation for 20 min, the concentrate and tailings were collected, filtered, dried and subjected to analysis for calculating iron recovery. 3. Results and discussion 3.1. Selective reduction roasting of ZLR 3.1.1. The effect of reduction roasting conditions on zinc ferrite decomposition During the reduction roasting process, a series of possible reactions occurred are listed as follows:
3ZnFe2 O4 þ CO ¼ 3ZnO þ 2Fe3 O4 þ CO2
Fig. 3. Effect of experimental conditions on the decomposition of zinc ferrite.
ð3-1Þ
106
H. Yan et al. / Minerals Engineering 55 (2014) 103–110
Fig. 4. TG curves of the ZLR under different reduction roasting conditions.
ZnFe2 O4 þ CO ¼ ZnO þ 2FeO þ CO2
ð3-2Þ
ZnFe2 O4 þ FeO ¼ ZnO þ Fe3 O4
ð3-3Þ
Fe3 O4 þ CO ¼ FeO þ CO2
ð3-4Þ
The effects of CO concentration, CO/CO + CO2 ratio, roasting temperature and duration time on the decomposition of zinc ferrite are shown in Fig. 3. It can be seen from Fig. 3a that the content of zinc oxide increased from 5.59% (unroasted samples) to 33.64% (2% CO), 60.63% (4% CO), 69.47% (6% CO), 76.61% (8% CO), 88.39% (10% CO), respectively. The increased zinc oxide comes from the decomposition of zinc ferrite and it is evident that the decomposition rate of zinc ferrite has a positive correlation with CO concentration. However, the Fe2+/TFe ratio increased sharply and exceeded 97.59% after roasted with 10% CO. The expected product Fe3O4 is easily over reduced to FeO and it is adverse to the separation of zinc and iron in the subsequent acid leaching process. Fig. 3b shows the influence of CO/CO + CO2 ratio on the decomposition of zinc ferrite. The Fe2+/TFe ratio decreased from 89.39% (CO/CO + CO2 ratio at 100%) to 46.39% (CO/CO + CO2 ratio at 15%) after roasted at 750 °C for 120 min with 8% CO. The results confirmed that CO2 can prevent Fe3O4 over reduced to FeO, namely, reactions (3-2)and (3-4)can be controlled in some extent. The effect of temperature on the decomposition of zinc ferrite was studied. The results in Fig. 3c show that the proportion of ferrous in total iron increased from 35.62% (roasted at 660 °C) to 48.88%
(roasted at 810 °C) and the content of zinc oxide increased from 27.98% (roasted at 660 °C) to 64.85% (roasted at 810 °C). Fig. 3d indicates the decomposition rate of zinc ferrite has a positive correlation with roasting time in the first 120 min. After that, the content of zinc oxide and proportion of ferrous in total iron remains at the stable level. In conclusion, the optimum reduction roasting operations should be performed at 750 °C for 90 min with 8% CO, and CO/CO + CO2 ratio at 30%.
3.1.2. Reduction behavior and mineralogical variation TG curves of the ZLR under different reduction conditions are shown in Fig. 4. Fig. 4a shows that the weight fraction of the leaching residue decreased gradually with the extension of duration time, and the weight loses faster with higher CO concentration. The results indicate that CO concentration exerts a positive effect on the reduction of zinc ferrite in ZLR. After roasted with different CO concentration at 750 °C for 120 min, the maximum weight loss are 2.33% (2% CO), 3.95% (4% CO), 5.09% (6% CO) , 5.83% (8% CO), 6.66% (10% CO. Fig. 4b shows that there is a fast weight-losing stage at the early 60 min and platforms appear gradually with the CO/CO + CO2 ratio less than 30%. In addition, the weight-losing rate increased with the increase of CO/CO + CO2 ratio and there is no platform appears when CO/CO + CO2 ratio over 40%. The results indicate that in the first 60 min, zinc ferrite decomposed effectively and CO/CO + CO2 ratio do not matter the reaction basically. After 60 min, platforms appeared gradually and present a negative
Fig. 5. X-ray diffraction pattern of the reduced ZLR with different CO concentration.
H. Yan et al. / Minerals Engineering 55 (2014) 103–110
107
Fig. 6. X-ray diffraction pattern of the reduced ZLR with different CO/CO + CO2 ratio.
Fig. 7. The SEM and EDS images of the reduced ZLR processed with 8% CO, 30% CO/CO + CO2 at 750 °C for 90 min.
correlation between CO/CO + CO2 ratio and weight losing rate. The maximum weight loss of zinc leaching residue with CO/CO + CO2 ratio range from 15% to 100% at 750 °C for 120 min are 3.60%(15%), 3.67%(20%), 3.88%(25%), 4.0%(30%), 4.5%(40%), 4.84%(50%) and 5.83%(100%), respectively. The XRD results of reduced ZLR under various CO concentration and CO/CO + CO2 ratio are shown in Figs. 5 and 6. The peaks
intensity and position indicate that the reduced ZLR is mostly composed of zinc oxide (ZnO), magnetite (Fe3O4), zinc ferrite (ZnFe2O4), zinc sulfide (ZnS) and willemite (Zn2SiO4), but the intensity and position of their characteristic peaks changed in different reduction roasting conditions. Fig. 5 shows that the characteristic peaks of zinc oxide (ZnO) appeared after reduced with CO concentration over 2% and its intensity presented a rising tendency with the
108
H. Yan et al. / Minerals Engineering 55 (2014) 103–110
Fig. 8. Effect of leaching conditions on the remove rate of iron and zinc.
Fig. 9. The SEM and EDS images of the residue leached with 90 g/lH2SO4, 10:1 liquid solid ratio, at 35 °C for 60 min.
increase of CO concentration. What is more, the peak position of zinc ferrite (ZnFe2O4) moved to magnetite (Fe3O4) gradually during the reduction process. The phenomenon can demonstrate the
selective decomposition of zinc ferrite. However, zinc ferrite is over reduced to ferrous oxide after reduced with CO concentration over 4%. The ferrous oxide will be leached out together with zinc oxide
H. Yan et al. / Minerals Engineering 55 (2014) 103–110
Fig. 10. Magnetic hysteresis loop of initial ZLR and iron concentrate.
Table 3 Iron content in different samples and iron recovery rate (%). Initial residue
Leached residue
Iron concentrate
Tailing
Iron recovery rate
23.91
37.14
43.01
14.73
80.90
(ZnO) which makes it difficult to separate iron from zinc. So a certain amount of CO2 was added to the reduction gases and the influence of CO/CO + CO2 ratio on the reduction products are shown in Fig. 6. Results show that the generation of ferrous oxide could be avoided effectively when the ratio of CO/CO + CO2 was maintained less than 30%. Scanning Electron Microscopy associated with Energy Dispersive Spectroscopy (SEM / EDS) techniques were used for the morphology study of the reduced ZLR (roasted at 750 °C with 8% CO and 30% CO/CO + CO2 for 90 min) and the results are shown in Fig. 7. It is seen that the reduced sample appeared in obvious stratification and one point in each layer was selected to analysis the elements composition by the EDS. Fig. 7a (point in the innermost layer) shows Si, Pb, Mn, Al, Fe, K and some other metal elements are mutually embedded in the reduced residue. Fig. 7b reveals that zinc and iron still coexist in the interlayer. Further treatment is necessary to recover the remained zinc ferrite and Si, Pb, Mn, Al, Fe, K metals. From Fig. 7c (point in the outmost layer), it can be found that there are iron oxides generated after the reduction processing. 3.2. Acid leaching of the reduced ZLR 3.2.1. The effect of leaching conditions on zinc recovery During the acid leaching process, the following chemical reactions are assumed (Trung et al., 2011):
ZnO þ H2 SO4 ¼ ZnSO4 þ H2 O
ð3-5Þ
ZnFe2 O4 þ H2 SO4 ¼ ZnSO4 þ Fe2 ðSO4 Þ3 þ 4H2 O
ð3-6Þ
ZnFe2 O4 þ H2 SO4 ¼ ZnSO4 þ Fe2 O3 þ H2 O
ð3-7Þ
ZnFe2 O4 þ H2 SO4 þ H2 O ¼ ZnSO4 þ 2FeðOHÞ3
ð3-8Þ
The effects of leaching time, leaching temperature, sulfuric acidity and liquid solid ratio on the remove rate of zinc and iron are shown in Fig. 8. It could be seen from Fig. 8a that with the extension of leaching time, the remove rate of zinc and iron
109
increased sharply at the first 15 min. However, the remove rate of zinc and iron increased slightly from 20 min to120 min due to the increase of path length for the diffusion of ions caused by the increasing of ions’ concentration in solutions and the reduction of reaction surface. Fig. 8b reveals that the remove rate of zinc and iron was significantly affected by the increase of leaching temperature. With the increase of temperature, the leaching efficiency of zinc increased sharply and exceeded 71.44% at 75 °C. But, 41.86% of iron was leached out which will adversely affect the separation of zinc and iron. This is due to the high temperature can increase the solubility of zinc ferrite: promote the reaction of Eqs. (3-6)–(3-8). As results shown in Fig. 8c, the remove rate of zinc increased from 57.90% (50 g/l) to 62.08% (130 g/l) gradually with the increase of sulfuric acidity. The remove rate of iron changed from15.96% (50 g/l) to 28.77% (130 g/l). The results indicate that sulfuric acidity exerts little influence on the zinc recovery. The results illustrated in Fig. 8d show the increase of liquid solid ratio has only limited effect on the remove rate when it is over 10:1 for the sulfuric acid is more than stoichiometric amount to dissolve zinc oxide. To achieve the selective leaching of zinc and to avoid lengthy iron precipitation process in the purification of the leaching solution, the remove rate of iron should be maintained at a low level on condition that the remove rate of zinc is high. According to the findings, the acid leaching operation performed at 35 °C for 60 min with 90 g/L H2SO4 and liquid/solid ratio at 10:1 can make the zinc remove rate reaches 61.38% while iron remove rate is only 25.34%. 3.2.2. The morphology of leached residue Scanning Electron Microscopy associated with Energy Dispersive Spectroscopy (SEM/EDS) techniques were used for the morphology study of residue after sulfuric acid leaching and the results are shown in Fig. 9. Three typical examples of SEM micrographs were selected to analysis the elements distribution in the leached residue. From Fig. 9a, it is seen that the particles of iron oxides remained in the residue for which may be embedded in impurities during acid leaching. The unoxidized zinc sulfide was found in Fig. 9b and this is one of the main reasons for low zinc leaching rate. Fig. 9c presents a grain contain Si, Mn, Mg, Fe, Ca and some other metal elements and it is similar with Fig. 7a. 3.3. Magnetic separation of the leached residue Magnetic separation was used to recover iron from the products of reduction roasting and acid leaching. The magnetic hysteresis loop of the initial ZLR and magnetic concentrate were investigated by Vibrating Sample Magnetometer and the results were shown in Fig. 10. The curves show that the Ms varied from 2.40 emu/g (initial residue) to 50.82 emu/g (iron concentrate) which indicate the realization of decomposition of zinc ferrite to magnetic iron by selective reduction. Potassium dichromate titration and EDTA titration methods are used to quantitative analysis the content of iron in the initial ZLR, leached residue and magnetic separation products with the results shown in Table 3. Grade of iron increased from 23.99% to 43.01% after a series of processing (ball-mining at 300 r/min for 15 min with ball to material ratio at 10:1, and magnetic separation at 1160 G for 20 min) and its recovery was 80.9%. 4. Conclusions The investigations revealed the preferential conditions and limitations of simultaneously recover zinc and iron from zinc leaching residue by the combination of selective reduction roasting, acid leaching and magnetic separation process. Experiments found that zinc ferrite in ZLR can be effectively decomposed to ZnO and Fe3O4
110
H. Yan et al. / Minerals Engineering 55 (2014) 103–110
when reduction performed under 8% CO with 30% CO/CO + CO2 at 750 °C for 90 min. During the reduction process, the over reduction of iron to ferrous could be controlled by CO2 effectively. Under the optimum leaching conditions (leaching at 35 °C for 60 min with 90 g/l sulfuric acid and liquid/solid ratio at 10:1), the leaching rate of zinc was 61.38% while the leaching rate of iron was only 25.34%. After magnetic separate on the content of iron in concentrate reached 43.01% and the iron recovery reached 80.90%. Zinc and iron were embedded in the unoxidized zinc sulfide and impurities (Si, Ca, Pb) (see Fig. 9) leads to the low recovery rate. As a result, desulfurization is a significant issue on simultaneous recovery of zinc and iron, which will be carried out in the future work. Acknowledgements The authors gratefully acknowledge the teachers and senior fellow apprentices for their help of experiments. The authors would also like to thank the National High Technology Research and Development Program of China (2011AA061001), National Scientific Research Project of Welfare (Environmental) Industry (2011467062), National Science Found for Distinguished Young Scholars of China (50925417), National Nature Science Foundation of China (50830301) for support of this work. References Antrekowitsch, J., 2001. Hydrometallurgically recovering zinc from electric arc furnace dusts. Journal of Minerals, Metals and Materials Society 53 (12), 26–28. Besße, A.V., Borulu, N., Çopur, M., Çolak, S., Ata, O.N., 2010. Optimization of dissolution of metals from Waelz sintering waste (WSW) by hydrochloric acid solutions. Chemical Engineering Journal 162 (2), 718–722. Çoruh, Semra, Nuri, Ergun Osman, 2010. Use of fly ash, phosphogypsum and red mud as a liner material for the disposal of hazardous zinc leach residue waste. Journal of Hazardous Materials 173 (1–3), 468–473. Dimitrios Filippou, G.P.D., 1997. Steady-state modeling of zinc-ferrite hot-acid leaching. Metallurgical and Materials Transactions B 28 (4), 701–711.
Dimitrova, R.I., Moldovanska, N., Bonev, I.K., Zˇivkovic, Zˇ., 2000. Oxidation of marmatite. Thermochimica Acta 362 (1–2), 145–151. Dutra, A.J.B., Paiva, P.R.P., Tavares, L.M., 2006. Alkaline leaching of zinc from electric arc furnace steel dust. Minerals Engineering 19 (5), 478–485. Graydon, J.W., Kirk, D.W., 1988. A Microscopic study of the transformation of sphalerite particles during the roasting of zinc concentrate. Metallurgical and Materials Transactions B 19 (1), 141–146. Hoang, J., Reuter, M.A., Matusewicz, R., Hughes, S., Piret, N., 2009. Top submerged lance direct zinc smelting. Minerals Engineering 22 (9–10), 742–751. Holloway, P.C., Etsell, T.H., Murland, A.L., 2007. Use of Secondary Additives to Control the Dissolution of Iron during Na2CO3 Roasting of La Oroya Zinc Ferrite. Metallurgical and Materials Transactions B 38 (5), 793–808. Jha, M.K., Kumar, V., Singh, R.J., 2001. Review of hydrometallurgical recovery of zinc from industrial wastes. Resources, Conservation and Recycling 33 (1), 1–22. Langová, Š., Riplová, J., Vallová, S., 2007. Atmospheric leaching of steel-making wastes and the precipitation of goethite from the ferric sulphate solution. Hydrometallurgy 87 (3–4), 157–162. Langová, Š., Leško, J., Maty´sek, D., 2009. Selective leaching of zinc from zinc ferrite with hydrochloric acid. Hydrometallurgy 95 (3–4), 179–182. Peng, N., Peng, B., Chai, L.-Y., Li, M., Wang, J.-M., Yan, H., Yuan, Y., 2012. Recovery of iron from zinc calcines by reduction roasting and magnetic separation. Minerals Engineering 35, 57–60. Ramazan Orhan, M.E., Soner Altundog˘an, H., Özer, Ahmet, Tümen, Fikret, 1998. Heavy metal pollution potential of zinc leach residues discarded in Cinkur Plant. Engineering and Environmental Science 22 (3), 167–177. Trung, Z.H., Kukurugya, F., Takacova, Z., Orac, D., Laubertova, M., Miskufova, A., Havlik, T., 2011. Acidic leaching both of zinc and iron from basic oxygen furnace sludge. Journal of Hazardous Materials 192 (3), 1100–1107. Turan, M.D., Altundog˘an, H.S., Tümen, F., 2004. Recovery of zinc and lead from zinc plant residue. Hydrometallurgy 75 (1–4), 169–176. Vahidi, E., Rashchi, F., Moradkhani, D., 2009. Recovery of zinc from an industrial zinc leach residue by solvent extraction using D2EHPA. Minerals Engineering 22 (2), 204–206. Xia, D.K., Picklesi, C.A., 2000. Microwave caustic leaching of electric arc furnace dust. Minerals Engineering 13 (1), 79–94. Zhang, H.B., 1992. Chemical Phase Analysis of Ore and Industrial Product. Metallurgical Industry Press. Zhang, Y., Yu, X., Li, X., 2011. Zinc recovery from franklinite by sulphation roasting. Hydrometallurgy 109 (3–4), 211–214. Zhao, Y.C., Stanforth, R., 2000. Integrated hydrometallurgical process for production of zinc from electric arc furnace dust in alkaline medium. Journal of Hazardous Materials 80 (1–3), 223–240.
Contents lists available at ScienceDirect
Minerals Engineering journal homepage: www.elsevier.com/locate/mineng
A novel method to recover zinc and iron from zinc leaching residue Huan Yan a, Li-yuan Chai a,b,⇑, Bing Peng a,b, Mi Li a, Ning Peng a, Dong-ke Hou a a b
Institute of Environmental Science and Engineering, School of Metallurgy and Environment, Central South University, 410083 Changsha, Hunan, China Chinese National Engineering Research Center for Control & Treatment of Heavy Metal Pollution, 410083 Changsha, Hunan, China
a r t i c l e
i n f o
Article history: Received 18 June 2013 Accepted 26 September 2013 Available online 24 October 2013 Keywords: Zinc leaching residue Zinc ferrite Reduction roasting Acid leaching Magnetic separation
a b s t r a c t A novel method to recover zinc and iron from zinc leaching residue (ZLR) by the combination of reduction roasting, acid leaching and magnetic separation was proposed. Zinc ferrite in the ZLR was selectively transformed to ZnO and Fe3O4 under CO, CO2 and Ar atmosphere. Subsequently, acid leaching was carried out to dissolve zinc from reduced ZLR while iron was left in the residue and recovered by magnetic separation. The mineralogical changes of ZLR during the processes were characterized by XRF, TG, XRD, SEM–EDS and VSM. The effects of roasting and leaching conditions were investigated with the optimum conditions obtained as follows: roasted at 750 °C for 90 min with 8% CO and CO/CO + CO2 ratio at 30%; leached at 35 °C for 60 min with 90 g/l sulfuric acid and liquid to solid ratio at 10:1. The iron was recovered by magnetic separation with magnetic intensity at 1160 G for 20 min. Under the optimum operation, 61.38% of zinc was recovered and 80.9% of iron recovery was achieved. This novel method not only realized the simultaneous recovery of zinc and iron but also solved the environmental problem caused by the storage of massive ZLR. Ó 2013 Elsevier Ltd. All rights reserved.
1. Introduction Currently, more than 85% of the world’s metal zinc is produced by the conventional hydrometallurgical approach including oxidative roasting, acid leaching, purification, and electrowinning processes (RLPE) (Jha et al., 2001; Turan et al., 2004). During the oxidative roasting process, ZnS is converted to ZnO, but a significant fraction reacts with the iron impurities to form zinc ferrite (Dimitrova et al., 2000; Langová et al., 2009; Peng et al., 2012; Vahidi et al., 2009). Zinc ferrite is hardly soluble in mild acidic conditions, and a huge quantity of leaching residue is produced in the subsequent leaching process. The residues could be used to recover metallic values or it may be disposed off (Soner Altundog˘an et al., 1998). Pyrometallurgical and hydrometallurgical processes have been studied to recycle zinc and iron from ZLR. The most typical pyrometallurgical method is Waelze process, and the recovery of zinc could reach to 90–95% in this process. However, Waelze process consumes huge energy to maintain the high reaction temperature (above 1300 K), and large amount of high iron-bearing secondary residue is generated (Besße et al., 2010). These shortcomings also exist in other pyrometallurgical methods including Ausmelt technology (Hoang et al., 2009), roasting with Na2CO3 (Holloway
⇑ Corresponding author at: Institute of Environmental Science and Engineering, School of Metallurgy and Environment, Central South University, 410083 Changsha, Hunan, China. E-mail address: [email protected] (L.-y. Chai). 0892-6875/$ - see front matter Ó 2013 Elsevier Ltd. All rights reserved. http://dx.doi.org/10.1016/j.mineng.2013.09.015
et al., 2007), roasting with Na2S (Zhang et al., 2011), and some other methods (Antrekowitsch, 2001; Besße et al., 2010; Çoruh and Nuri, 2010). The most common hydrometallurgical process is to recycle zinc and iron from ZLR in a hot concentrated sulfuric acid followed by precipitation of the dissolved iron values from solution as jarosite, goethite or hematite (Graydon, 1988). High leaching rate both of zinc and iron could be obtained in these processes, but the removal of iron and other unwanted metals from the leaching solution is difficult (Dimitrios Filippou, 1997). Hydrochloric acid leaching (Langová et al., 2009), alkaline solution leaching (Dutra et al., 2006; Zhao and Stanforth, 2000), microwave caustic leaching (Xia and Picklesi, 2000), D2EHPA leaching (Vahidi et al., 2009), and leaching with various solvents have been studied (Langová et al., 2007). These hydrometallurgical processes are more economical because of lower capital and operating costs, but the purification process becomes more complex. Therefore finding a cost-effective and environmental-friendly process to recover zinc and iron from zinc leaching residue remains the major challenge. In this study, a simultaneously recycling zinc and iron from ZLR by the combination of selective reduction, acid leaching and magnetic separation is investigated. The objective of this recycling technology is to decompose ZnFe2O4 to ZnO and Fe3O4 selectively. Then, ZnO and Fe3O4 could be separated by acid leaching and magnetic separation, respectively. This study focus on the various operating parameters in the three stages under laboratory conditions to control phase transformation and search the optimum experiments conditions. The phase transformation, morphology variation and magnetism changes of the samples are detected by
104
H. Yan et al. / Minerals Engineering 55 (2014) 103–110
XRD, SEM and VSM to interpret some phenomenon observed in experiments and the decomposition procedure of ZnFe2O4.
500
2. Experimental approach
400
2 1
1----ZnFe2O4
6----Zn2SiO4
2----Fe3O4
7----PbS 8----Pb4SiO6
2.1. Materials and analysis ZLR used in this study was obtained from a zinc hydrometallurgical plant in Hunan province, China. The samples were dried at 105 °C for 24 h before grinding and sieved to obtain required fractions. Elements composition of the residue was determined by XRF and results in Table 1 show that the main contents in ZLR used in this study are Zn-19.57%, Fe-23.91%, S-6.40%, Si-4.47%, Pb-4.35%, Ca-2.21%. The crystalline phases of the samples were investigated by X-ray powder diffraction (Rigaku, TTR-III) in 2h scale using Cu Ka radiation (k = 1.5406 Å, 50 kV and 100 mA) at the scanning rate of 10°/min vary from 10° to 80°. As Fig. 1 shows that zinc ferrite (ZnFe2O4), magnetite (Fe3O4), zinc sulfide (ZnS), zinc silicate (Zn2SiO4), and calcium sulfate (CaSO4) are the main crystal mineral phases in the residue. The morphological changes were detected by Scanning Electron Microscopy (SEM, JEOL.LTD, JSM-6360LV). Vibration sample magnetometer (VSM) (Model: HH-15) was applied to assess the magnetism of magnetic concentrates. Potassium dichromate titration and EDTA titration methods were used to quantitative analysis the proportion of ferrous in total iron and content of zinc oxide to evaluate the effect of experiments (Zhang, 1992). 2.2. Experimental procedure 2.2.1. Reduction roasting The experiments were performed according to the flow sheet as shown in Fig. 2. Reduction roasting process was firstly operated in thermo-gravimetric apparatus (STA449F3) to decompose zinc ferrite to ZnO and Fe3O4 selectively. The ZLR was flatted on the inner wall of the silica crucible and pick up the suction. Then, the sample was heat up to determined temperature at 10 K/min under 100 ml/min Ar. As soon as the temperature reached, a proportion fraction of reduction gases (CO + CO2 + Ar) was added to take place
Intensity (counts)
3----CaSO4
9----FeS
4----PbSO4 5----ZnS
300 200 7
2 1
3
100
2 1
0 10
6 5 9 4 4 74 5
20
2 1
2 1 4
30
40
2 1
2 1 2 65 8 1
2 2 1 1
50
60
70
80
Angle (2-Theta deg.) Fig. 1. X-ray diffraction pattern of ZLR.
Table 1 Chemical composition of ZLR by XRF. Elements
Content (wt.%)
Elements
Content (wt.%)
Element
Content (wt.%)
Fe Zn S Si Pb Ca Mn Al
23.91 19.57 6.40 4.47 4.35 2.21 1.53 1.15
Cu As Cd Mg K Ba Sr Sn
0.8 0.52 0.31 0.26 0.24 0.20 0.16 0.16
Cl Ti In P Ag Cr Ni Mo
0.067 0.057 0.054 0.043 0.037 0.018 0.011 0.010
of Ar at 100 ml/min, respectively. After reduction roasting, the mixed reduction gases (CO + CO2 + Ar) were changed to Ar again to cool down the samples. The reduced products were stored in a tightly closed jar waiting for analysis. The influences of reduction roasting parameters such as CO concentration, CO/CO + CO2 ratio, roasting temperature and duration time were studied as detailed shown in Table 2. The content of zinc
Fig. 2. Flow sheet of the process of simultaneously recovering zinc and iron from ZLR.
105
H. Yan et al. / Minerals Engineering 55 (2014) 103–110 Table 2 Conditions used in each operation. Operations
Parameters
a
b
c
d
Reduction roasting
CO (%) CO/CO + CO2 (%) T (°C) t (min)
2–10 100 750 120
8 15–50 750 120
8 20 660–810 120
8 20 750 30–180
Acid leaching
T (°C) Acidity (g/L) S/L t (min)
25–65 90 9:1 30
35 70–120 9:1 30
35 90 5:1–13:1 30
35 90 9:1 10–60
Magnetic separation
t (min) Current intensity (G)
20 1160
– –
– –
– –
oxide and Fe2+/TFe ratio in reduced products were used to evaluate the reduction roasting effects. 2.2.2. Acid leaching Sulfuric acid leaching was conducted to selectively separate ZnO from the reduced ZLR. The leaching process was carried out in an agitator with which the leaching temperature and the speed of agitation can be controlled. The range of acidity, leaching temperature, leaching time and liquid to solid ratio in each operation were listed in Table 2. The remove rate of zinc and iron was used to evaluate the separation rate of zinc and iron. 2.2.3. Magnetic separation Magnetic separation was performed on a slurry of leaching residue (10% in solid) using a Davis tube magnetic separator (XCGS). Firstly, a certain amount of samples were milled with ball material
ratio at 10:1 and rotational speed at 300 r/min for 15 min to disperse the particles. Then, the magnetic separation was carried out to recover iron with magnetic intensity at 1160 G. After magnetic separation for 20 min, the concentrate and tailings were collected, filtered, dried and subjected to analysis for calculating iron recovery. 3. Results and discussion 3.1. Selective reduction roasting of ZLR 3.1.1. The effect of reduction roasting conditions on zinc ferrite decomposition During the reduction roasting process, a series of possible reactions occurred are listed as follows:
3ZnFe2 O4 þ CO ¼ 3ZnO þ 2Fe3 O4 þ CO2
Fig. 3. Effect of experimental conditions on the decomposition of zinc ferrite.
ð3-1Þ
106
H. Yan et al. / Minerals Engineering 55 (2014) 103–110
Fig. 4. TG curves of the ZLR under different reduction roasting conditions.
ZnFe2 O4 þ CO ¼ ZnO þ 2FeO þ CO2
ð3-2Þ
ZnFe2 O4 þ FeO ¼ ZnO þ Fe3 O4
ð3-3Þ
Fe3 O4 þ CO ¼ FeO þ CO2
ð3-4Þ
The effects of CO concentration, CO/CO + CO2 ratio, roasting temperature and duration time on the decomposition of zinc ferrite are shown in Fig. 3. It can be seen from Fig. 3a that the content of zinc oxide increased from 5.59% (unroasted samples) to 33.64% (2% CO), 60.63% (4% CO), 69.47% (6% CO), 76.61% (8% CO), 88.39% (10% CO), respectively. The increased zinc oxide comes from the decomposition of zinc ferrite and it is evident that the decomposition rate of zinc ferrite has a positive correlation with CO concentration. However, the Fe2+/TFe ratio increased sharply and exceeded 97.59% after roasted with 10% CO. The expected product Fe3O4 is easily over reduced to FeO and it is adverse to the separation of zinc and iron in the subsequent acid leaching process. Fig. 3b shows the influence of CO/CO + CO2 ratio on the decomposition of zinc ferrite. The Fe2+/TFe ratio decreased from 89.39% (CO/CO + CO2 ratio at 100%) to 46.39% (CO/CO + CO2 ratio at 15%) after roasted at 750 °C for 120 min with 8% CO. The results confirmed that CO2 can prevent Fe3O4 over reduced to FeO, namely, reactions (3-2)and (3-4)can be controlled in some extent. The effect of temperature on the decomposition of zinc ferrite was studied. The results in Fig. 3c show that the proportion of ferrous in total iron increased from 35.62% (roasted at 660 °C) to 48.88%
(roasted at 810 °C) and the content of zinc oxide increased from 27.98% (roasted at 660 °C) to 64.85% (roasted at 810 °C). Fig. 3d indicates the decomposition rate of zinc ferrite has a positive correlation with roasting time in the first 120 min. After that, the content of zinc oxide and proportion of ferrous in total iron remains at the stable level. In conclusion, the optimum reduction roasting operations should be performed at 750 °C for 90 min with 8% CO, and CO/CO + CO2 ratio at 30%.
3.1.2. Reduction behavior and mineralogical variation TG curves of the ZLR under different reduction conditions are shown in Fig. 4. Fig. 4a shows that the weight fraction of the leaching residue decreased gradually with the extension of duration time, and the weight loses faster with higher CO concentration. The results indicate that CO concentration exerts a positive effect on the reduction of zinc ferrite in ZLR. After roasted with different CO concentration at 750 °C for 120 min, the maximum weight loss are 2.33% (2% CO), 3.95% (4% CO), 5.09% (6% CO) , 5.83% (8% CO), 6.66% (10% CO. Fig. 4b shows that there is a fast weight-losing stage at the early 60 min and platforms appear gradually with the CO/CO + CO2 ratio less than 30%. In addition, the weight-losing rate increased with the increase of CO/CO + CO2 ratio and there is no platform appears when CO/CO + CO2 ratio over 40%. The results indicate that in the first 60 min, zinc ferrite decomposed effectively and CO/CO + CO2 ratio do not matter the reaction basically. After 60 min, platforms appeared gradually and present a negative
Fig. 5. X-ray diffraction pattern of the reduced ZLR with different CO concentration.
H. Yan et al. / Minerals Engineering 55 (2014) 103–110
107
Fig. 6. X-ray diffraction pattern of the reduced ZLR with different CO/CO + CO2 ratio.
Fig. 7. The SEM and EDS images of the reduced ZLR processed with 8% CO, 30% CO/CO + CO2 at 750 °C for 90 min.
correlation between CO/CO + CO2 ratio and weight losing rate. The maximum weight loss of zinc leaching residue with CO/CO + CO2 ratio range from 15% to 100% at 750 °C for 120 min are 3.60%(15%), 3.67%(20%), 3.88%(25%), 4.0%(30%), 4.5%(40%), 4.84%(50%) and 5.83%(100%), respectively. The XRD results of reduced ZLR under various CO concentration and CO/CO + CO2 ratio are shown in Figs. 5 and 6. The peaks
intensity and position indicate that the reduced ZLR is mostly composed of zinc oxide (ZnO), magnetite (Fe3O4), zinc ferrite (ZnFe2O4), zinc sulfide (ZnS) and willemite (Zn2SiO4), but the intensity and position of their characteristic peaks changed in different reduction roasting conditions. Fig. 5 shows that the characteristic peaks of zinc oxide (ZnO) appeared after reduced with CO concentration over 2% and its intensity presented a rising tendency with the
108
H. Yan et al. / Minerals Engineering 55 (2014) 103–110
Fig. 8. Effect of leaching conditions on the remove rate of iron and zinc.
Fig. 9. The SEM and EDS images of the residue leached with 90 g/lH2SO4, 10:1 liquid solid ratio, at 35 °C for 60 min.
increase of CO concentration. What is more, the peak position of zinc ferrite (ZnFe2O4) moved to magnetite (Fe3O4) gradually during the reduction process. The phenomenon can demonstrate the
selective decomposition of zinc ferrite. However, zinc ferrite is over reduced to ferrous oxide after reduced with CO concentration over 4%. The ferrous oxide will be leached out together with zinc oxide
H. Yan et al. / Minerals Engineering 55 (2014) 103–110
Fig. 10. Magnetic hysteresis loop of initial ZLR and iron concentrate.
Table 3 Iron content in different samples and iron recovery rate (%). Initial residue
Leached residue
Iron concentrate
Tailing
Iron recovery rate
23.91
37.14
43.01
14.73
80.90
(ZnO) which makes it difficult to separate iron from zinc. So a certain amount of CO2 was added to the reduction gases and the influence of CO/CO + CO2 ratio on the reduction products are shown in Fig. 6. Results show that the generation of ferrous oxide could be avoided effectively when the ratio of CO/CO + CO2 was maintained less than 30%. Scanning Electron Microscopy associated with Energy Dispersive Spectroscopy (SEM / EDS) techniques were used for the morphology study of the reduced ZLR (roasted at 750 °C with 8% CO and 30% CO/CO + CO2 for 90 min) and the results are shown in Fig. 7. It is seen that the reduced sample appeared in obvious stratification and one point in each layer was selected to analysis the elements composition by the EDS. Fig. 7a (point in the innermost layer) shows Si, Pb, Mn, Al, Fe, K and some other metal elements are mutually embedded in the reduced residue. Fig. 7b reveals that zinc and iron still coexist in the interlayer. Further treatment is necessary to recover the remained zinc ferrite and Si, Pb, Mn, Al, Fe, K metals. From Fig. 7c (point in the outmost layer), it can be found that there are iron oxides generated after the reduction processing. 3.2. Acid leaching of the reduced ZLR 3.2.1. The effect of leaching conditions on zinc recovery During the acid leaching process, the following chemical reactions are assumed (Trung et al., 2011):
ZnO þ H2 SO4 ¼ ZnSO4 þ H2 O
ð3-5Þ
ZnFe2 O4 þ H2 SO4 ¼ ZnSO4 þ Fe2 ðSO4 Þ3 þ 4H2 O
ð3-6Þ
ZnFe2 O4 þ H2 SO4 ¼ ZnSO4 þ Fe2 O3 þ H2 O
ð3-7Þ
ZnFe2 O4 þ H2 SO4 þ H2 O ¼ ZnSO4 þ 2FeðOHÞ3
ð3-8Þ
The effects of leaching time, leaching temperature, sulfuric acidity and liquid solid ratio on the remove rate of zinc and iron are shown in Fig. 8. It could be seen from Fig. 8a that with the extension of leaching time, the remove rate of zinc and iron
109
increased sharply at the first 15 min. However, the remove rate of zinc and iron increased slightly from 20 min to120 min due to the increase of path length for the diffusion of ions caused by the increasing of ions’ concentration in solutions and the reduction of reaction surface. Fig. 8b reveals that the remove rate of zinc and iron was significantly affected by the increase of leaching temperature. With the increase of temperature, the leaching efficiency of zinc increased sharply and exceeded 71.44% at 75 °C. But, 41.86% of iron was leached out which will adversely affect the separation of zinc and iron. This is due to the high temperature can increase the solubility of zinc ferrite: promote the reaction of Eqs. (3-6)–(3-8). As results shown in Fig. 8c, the remove rate of zinc increased from 57.90% (50 g/l) to 62.08% (130 g/l) gradually with the increase of sulfuric acidity. The remove rate of iron changed from15.96% (50 g/l) to 28.77% (130 g/l). The results indicate that sulfuric acidity exerts little influence on the zinc recovery. The results illustrated in Fig. 8d show the increase of liquid solid ratio has only limited effect on the remove rate when it is over 10:1 for the sulfuric acid is more than stoichiometric amount to dissolve zinc oxide. To achieve the selective leaching of zinc and to avoid lengthy iron precipitation process in the purification of the leaching solution, the remove rate of iron should be maintained at a low level on condition that the remove rate of zinc is high. According to the findings, the acid leaching operation performed at 35 °C for 60 min with 90 g/L H2SO4 and liquid/solid ratio at 10:1 can make the zinc remove rate reaches 61.38% while iron remove rate is only 25.34%. 3.2.2. The morphology of leached residue Scanning Electron Microscopy associated with Energy Dispersive Spectroscopy (SEM/EDS) techniques were used for the morphology study of residue after sulfuric acid leaching and the results are shown in Fig. 9. Three typical examples of SEM micrographs were selected to analysis the elements distribution in the leached residue. From Fig. 9a, it is seen that the particles of iron oxides remained in the residue for which may be embedded in impurities during acid leaching. The unoxidized zinc sulfide was found in Fig. 9b and this is one of the main reasons for low zinc leaching rate. Fig. 9c presents a grain contain Si, Mn, Mg, Fe, Ca and some other metal elements and it is similar with Fig. 7a. 3.3. Magnetic separation of the leached residue Magnetic separation was used to recover iron from the products of reduction roasting and acid leaching. The magnetic hysteresis loop of the initial ZLR and magnetic concentrate were investigated by Vibrating Sample Magnetometer and the results were shown in Fig. 10. The curves show that the Ms varied from 2.40 emu/g (initial residue) to 50.82 emu/g (iron concentrate) which indicate the realization of decomposition of zinc ferrite to magnetic iron by selective reduction. Potassium dichromate titration and EDTA titration methods are used to quantitative analysis the content of iron in the initial ZLR, leached residue and magnetic separation products with the results shown in Table 3. Grade of iron increased from 23.99% to 43.01% after a series of processing (ball-mining at 300 r/min for 15 min with ball to material ratio at 10:1, and magnetic separation at 1160 G for 20 min) and its recovery was 80.9%. 4. Conclusions The investigations revealed the preferential conditions and limitations of simultaneously recover zinc and iron from zinc leaching residue by the combination of selective reduction roasting, acid leaching and magnetic separation process. Experiments found that zinc ferrite in ZLR can be effectively decomposed to ZnO and Fe3O4
110
H. Yan et al. / Minerals Engineering 55 (2014) 103–110
when reduction performed under 8% CO with 30% CO/CO + CO2 at 750 °C for 90 min. During the reduction process, the over reduction of iron to ferrous could be controlled by CO2 effectively. Under the optimum leaching conditions (leaching at 35 °C for 60 min with 90 g/l sulfuric acid and liquid/solid ratio at 10:1), the leaching rate of zinc was 61.38% while the leaching rate of iron was only 25.34%. After magnetic separate on the content of iron in concentrate reached 43.01% and the iron recovery reached 80.90%. Zinc and iron were embedded in the unoxidized zinc sulfide and impurities (Si, Ca, Pb) (see Fig. 9) leads to the low recovery rate. As a result, desulfurization is a significant issue on simultaneous recovery of zinc and iron, which will be carried out in the future work. Acknowledgements The authors gratefully acknowledge the teachers and senior fellow apprentices for their help of experiments. The authors would also like to thank the National High Technology Research and Development Program of China (2011AA061001), National Scientific Research Project of Welfare (Environmental) Industry (2011467062), National Science Found for Distinguished Young Scholars of China (50925417), National Nature Science Foundation of China (50830301) for support of this work. References Antrekowitsch, J., 2001. Hydrometallurgically recovering zinc from electric arc furnace dusts. Journal of Minerals, Metals and Materials Society 53 (12), 26–28. Besße, A.V., Borulu, N., Çopur, M., Çolak, S., Ata, O.N., 2010. Optimization of dissolution of metals from Waelz sintering waste (WSW) by hydrochloric acid solutions. Chemical Engineering Journal 162 (2), 718–722. Çoruh, Semra, Nuri, Ergun Osman, 2010. Use of fly ash, phosphogypsum and red mud as a liner material for the disposal of hazardous zinc leach residue waste. Journal of Hazardous Materials 173 (1–3), 468–473. Dimitrios Filippou, G.P.D., 1997. Steady-state modeling of zinc-ferrite hot-acid leaching. Metallurgical and Materials Transactions B 28 (4), 701–711.
Dimitrova, R.I., Moldovanska, N., Bonev, I.K., Zˇivkovic, Zˇ., 2000. Oxidation of marmatite. Thermochimica Acta 362 (1–2), 145–151. Dutra, A.J.B., Paiva, P.R.P., Tavares, L.M., 2006. Alkaline leaching of zinc from electric arc furnace steel dust. Minerals Engineering 19 (5), 478–485. Graydon, J.W., Kirk, D.W., 1988. A Microscopic study of the transformation of sphalerite particles during the roasting of zinc concentrate. Metallurgical and Materials Transactions B 19 (1), 141–146. Hoang, J., Reuter, M.A., Matusewicz, R., Hughes, S., Piret, N., 2009. Top submerged lance direct zinc smelting. Minerals Engineering 22 (9–10), 742–751. Holloway, P.C., Etsell, T.H., Murland, A.L., 2007. Use of Secondary Additives to Control the Dissolution of Iron during Na2CO3 Roasting of La Oroya Zinc Ferrite. Metallurgical and Materials Transactions B 38 (5), 793–808. Jha, M.K., Kumar, V., Singh, R.J., 2001. Review of hydrometallurgical recovery of zinc from industrial wastes. Resources, Conservation and Recycling 33 (1), 1–22. Langová, Š., Riplová, J., Vallová, S., 2007. Atmospheric leaching of steel-making wastes and the precipitation of goethite from the ferric sulphate solution. Hydrometallurgy 87 (3–4), 157–162. Langová, Š., Leško, J., Maty´sek, D., 2009. Selective leaching of zinc from zinc ferrite with hydrochloric acid. Hydrometallurgy 95 (3–4), 179–182. Peng, N., Peng, B., Chai, L.-Y., Li, M., Wang, J.-M., Yan, H., Yuan, Y., 2012. Recovery of iron from zinc calcines by reduction roasting and magnetic separation. Minerals Engineering 35, 57–60. Ramazan Orhan, M.E., Soner Altundog˘an, H., Özer, Ahmet, Tümen, Fikret, 1998. Heavy metal pollution potential of zinc leach residues discarded in Cinkur Plant. Engineering and Environmental Science 22 (3), 167–177. Trung, Z.H., Kukurugya, F., Takacova, Z., Orac, D., Laubertova, M., Miskufova, A., Havlik, T., 2011. Acidic leaching both of zinc and iron from basic oxygen furnace sludge. Journal of Hazardous Materials 192 (3), 1100–1107. Turan, M.D., Altundog˘an, H.S., Tümen, F., 2004. Recovery of zinc and lead from zinc plant residue. Hydrometallurgy 75 (1–4), 169–176. Vahidi, E., Rashchi, F., Moradkhani, D., 2009. Recovery of zinc from an industrial zinc leach residue by solvent extraction using D2EHPA. Minerals Engineering 22 (2), 204–206. Xia, D.K., Picklesi, C.A., 2000. Microwave caustic leaching of electric arc furnace dust. Minerals Engineering 13 (1), 79–94. Zhang, H.B., 1992. Chemical Phase Analysis of Ore and Industrial Product. Metallurgical Industry Press. Zhang, Y., Yu, X., Li, X., 2011. Zinc recovery from franklinite by sulphation roasting. Hydrometallurgy 109 (3–4), 211–214. Zhao, Y.C., Stanforth, R., 2000. Integrated hydrometallurgical process for production of zinc from electric arc furnace dust in alkaline medium. Journal of Hazardous Materials 80 (1–3), 223–240.