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Sequential stepwise recovery of selected metals from flue dusts of secondary copper smelting
Journal of Cleaner Production xxx (2014) 1e8
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Sequential stepwise recovery of selected metals from flue dusts of secondary copper smelting Li Qiang a, Isabel S.S. Pinto b, Zhao Youcai a, * a b
The State Key Laboratory of Pollution Control and Resource Reuse, Tongji University, Shanghai 200092, China REQUIMTE, Faculty of Engineering, University of Porto, Rua Dr.Roberto Frias, 4200-465 Porto, Portugal
a r t i c l e i n f o
a b s t r a c t
Article history: Received 4 August 2013 Received in revised form 22 March 2014 Accepted 24 March 2014 Available online xxx
A novel and clean process was developed to convert flue dusts of secondary copper smelters into valueadded products. The process consists of three main steps: NaOH leaching, two-stage combined electrolysis, and deep purification of recycle solution. Initially, 80e92% Zn and Pb were dissolved in 5 M NaOH at 80 C, whereas Cu was concentrated in the residue. Yates’ algorithm was used to determine the main effects and interactions of the leaching factors. The leach solution was then electrolyzed at 100e250 A/m2 and high purity Pb (>97%) was separated from the alkaline solution. Subsequently, a pulsed current was introduced to obtain ultrafine zinc powders (w30 mm) with the best performance occurred at Ton (current-on) ¼ 15 ms, and Toff (current-off) ¼ 10 ms. Finally, Cl and Al were precipitated by evaporation-condensation and CaO addition, respectively. The combination of alkaline media and pulse current enables the green route without distilled water washing and surfactant addition, thus minimizing effluent emissions. This environmentally friendly method has promise to treat other industrial wastes due to its low cost and simple design. Ó 2014 Elsevier Ltd. All rights reserved.
Keywords: Green process Metal recovery Yates’ algorithm Ultrafine zinc powder High pure lead Hazardous wastes
1. Introduction Metallurgical industries generate vast quantities of different types of wastes such as electric arc furnace (EAF) dust, basic oxygen furnace (BOF) sludge, jarosite residue and flue dust of secondary copper smelting (Dvorak and Jandova, 2006; Ju et al., 2011). On the basis of rough statistics, no less than one million tons of these wastes have been produced in China per year (Li et al., 2011). These wastes could be used to recover metallic values or they may be disposed of. However, the disposal of such material is now becoming expensive due to increasingly stringent environmental regulations. Furthermore, the chemical nature of these dust/ash particles is such that these are classified as hazardous waste under the US Environmental Protection Agency classification. In view of the above, there has been an increasing interest in developing processes for the recovery of metals from these wastes. Usually, pyrometallurgical and hydrometallurgical processes are employed for treating such residues. The pyrometallurgical processes face the problems of generation of worthless residues and
* Corresponding author. Fax: þ86 21 65982684. E-mail addresses: [email protected], [email protected] ( Zhao Youcai).
costly equipment investment as well as quite high energy consumption (Pinto and Soares, 2012, 2013). Thus, much more attention has been paid to hydrometallurgical processes. For example, Dutra et al. (2006) extracted the zinc present in EAF dust with different alkaline leaching techniques. Oustadakis and coauthors developed a process of H2SO4 leaching, solvent extraction and electrowinning to recover zinc from EAF dust (Oustadakis et al., 2010; Tsakiridis et al., 2010). Langova et al. (2007) examined the sulfuric acid leaching of EAF dust and studied the influence of acid concentration, temperature, time, and liquid/solid ratio. Zinc extraction reached almost 100% and iron extraction exceeded 90% in 3 M H2SO4 at 80 C and liquid/solid ratio of 5 after 6 h. Furthermore, they found that a good selectivity with regard to zinc was achieved with 0.1e0.3 M H2SO4 at 80 C. Ju et al. (2011) proposed a clean hydrometallurgical route to recover zinc, silver, lead, copper, cadmium and iron from hazardous jarosite residues produced by zinc hydrometallurgy, whereas Trung et al. (2011) concentrated on the zinc leaching from fine-grain BOF sludge in acid medium. However, there is still not enough information about how to process wastes from the secondary Cu industry. The flue dusts used in this work were obtained from the city of Fuyang, which is situated in Zhejiang, a developed southeastern province in China; it is one of the biggest national centers for secondary Cu industries. The composition of the dusts varies
http://dx.doi.org/10.1016/j.jclepro.2014.03.085 0959-6526/Ó 2014 Elsevier Ltd. All rights reserved.
Please cite this article in press as: Qiang, L., et al., Sequential stepwise recovery of selected metals from flue dusts of secondary copper smelting, Journal of Cleaner Production (2014), http://dx.doi.org/10.1016/j.jclepro.2014.03.085
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considerably, and it is dependent not only on the waste used, but also on the operating conditions. However, some general trends have been noted. For example, most of these dusts are rich in Zn, Cu, Cl, Pb, and Al. It is advisable to leach the dusts with caustic soda, because the chlorides will not be tolerated in the acid electrolyte, as even a very small amount can cause severe corrosion problems and thereby damage the electrolysis (Gupta et al., 1989; Guresin and Topkaya, 1998). Moreover, the washing section for removing Cl produces secondary pollutants and complicates the process. During caustic soda process, however, the chloride concentration increased over 25 g/L without any trouble being observed (Frenay et al., 1986). In particular no chlorine evolution was noted in the electrolysis cells, and formation of hypochlorite would be expected from the following reaction:
Cl þ H2 O/HClO þ Hþ þ 2e :
(1)
In addition, most of copper (>90%) is left in the residue when using this alkaline leach reagent. Furthermore, in the alkaline zinc electrowinning process hydrogen evolution on the cathode surface is even more impeded than the one in acidic electrowinning. Current efficiencies (CE) of 97e99% can thus be reached when high current densities are applied (Gurmen and Emre, 2003; St-Pierre and Piron, 1990). This behavior promotes the generation of more profitable products with a lower energy cost. On the basis of the above analysis, a hydrometallurgical route is presented in this work, as shown in Fig. 1, for the metal recovery from the flue dusts of secondary copper smelters. In the alkaline zinc electrowinning, Na2S has been used to remove Pb from the leach solution, and this approach can recover lead selectively and quantitatively (Zhao and Stanforth, 2001). However, it is time consuming to separate Pb from alkaline solution, and Na2S is a toxic agent. Therefore, the present research focuses on the lead recovery from the alkaline media by low current densities electrolysis. In the conventional process, surfactants are widely applied to control the
size and morphology of zinc powders (Ghavami and Rafiei, 2006); however the accumulation of these additives could damage the recycle solution quality and increase effluent emissions. Hence, other potential methods are clearly needed to produce ultrafine zinc powders with uniform size. In this work, pulse electrolysis is employed to obtain zinc powder due to its variety of mass transport situations and electrocrystallization conditions. However, the majority of studies on pulse electrolysis have focused on its potential advantages in the electroplating industry (Chandrasekar et al., 2010; Saber et al., 2003; Youssef et al., 2008). Information is lacking regarding its use in powder production. This work describes a novel and efficient method to extract Pb from an alkaline leaching solution of the sample. Attempt was also made to produce uniform and fine zinc powders with pulse current. Subsequently, we reported the techniques for removing Cl and Al from the resultant alkaline electrolyte. 2. Material and methods 2.1. Characterization of the flue dusts The sample was dried and homogenized before the leaching experiments, and its chemical composition is listed in Table 1. The chemical analysis was determined by ICP-OES (AGILENT 720ES) and Ion Chromatography (ICS-1000). It demonstrates that the major elements present in the sample are zinc, copper, lead, aluminum and chlorine. An XRD (BRUKER D8 ADVANCE) analysis of the asreceived dust is displayed in Fig. 2. It shows that most of metal elements are in combination with oxygen, but there can also be chloride. 2.2. Computer chemical simulations Chemical speciation calculations were carried out using the MINEQLþ software (version 4.5) (Schecher and McAvoy, 2003), and the species distribution of metals at different pH values were determined. Metal speciation analysis with MINEQLþ generates chemical equilibrium concentrations of all species being considered in the model by the program reactions, based on component stability constants (Martell and Smith, 2004) and total molar metal concentrations. For computer simulations, total molar metal concentrations were calculated considering a solid to liquid (S/L) ratio of 100 g/L and assuming that the total amount of metals in the dust were in solution. The simulations allow predicting the optimum pH range to solubilize the target metals. 2.3. Procedure The proposed process consists of three main stages, including NaOH leaching, two-stage combined electrolysis, and deep purification of recycle solution. With this method, the stepwise extraction of Cu, Pb, Zn and Al was achieved. Moreover, ICP-OES, XRD, SEM-EDS (PHILIPS XL30), IC, particle size analyses (Malvern 3000), and electrochemical methods were used in this study. 2.3.1. Leaching The leaching processes were carried out at 30e90 C under atmospheric pressure. Specific amounts of flue dust samples were Table 1 Chemical analysis of the main elements present in the dust.
Fig. 1. General scheme for treating flue dusts from secondary copper smelters in this process.
Element
Zn
Cu
Pb
Al
Cl
Weight (%)
40.21
7.53
6.62
2.58
8.47
Please cite this article in press as: Qiang, L., et al., Sequential stepwise recovery of selected metals from flue dusts of secondary copper smelting, Journal of Cleaner Production (2014), http://dx.doi.org/10.1016/j.jclepro.2014.03.085
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Fig. 2. Mineralogical phases of flue dusts used.
added into a glass reactor containing a given volume of 3e6 M NaOH solution, and then the reactor rotated with an orbital shaker for 1e2 h (15 rpm). A series of leach tests were run, and the effects of variables on zinc recovery were assessed using Yates’ algorithm. Finally, the supernatant and the residues were analyzed by ICP-OES. 2.3.2. Electrolysis In electrowinning, stainless steel electrodes were used as both anode and cathode. Moreover, the anode to cathode distance was kept at 3 cm and all electrowinning experiments were carried out at room temperature (30e50 C). Current, cell voltage, and metal concentrations were recorded during the electrolysis, while at least two replicates were used for every measurement. Finally, the cathode CE was calculated from the mass gain obtained after electrolysis. Electrolysis of lead was attained by applying a low current density of (100e250 A/m2). At the end of electrowinning, Pb concentration in the final solution was quantified by ICP-OES to determine the efficiency of metal recovery. The purity of the product was also analyzed by ICP-OES and SEM-EDS. Zinc powders were produced by pulse direct current. The current-on time (Ton) and the current-off time (Toff) were optimized to obtain fine and homogeneous particles. The optimization was performed through a series of experiments in which Ton was varied from 5 to 25 ms and Toff from 5 to 20 ms at a constant pulse current density (1500e1800 A/m2). 2.4. Analysis method 2.4.1. Electrochemical measurements A CHI600D electrochemical workstation was used for the polarization studies, and all the polarization measurements were performed in a three-electrode Plexiglas cell. A couple of stainless steel plates were employed as working and auxiliary electrodes, while reference electrode was a saturated calomel electrode (SCE) with an appropriate lugging capillary. All potentials were recorded with respect to SCE. Cyclic voltammetries were carried out by initiating scans at 0.5 V versus SCE at a rate of 5 mV s1. 2.4.2. Powder examination SEM and XRD were used to characterize the surface morphology and the preferred crystal orientation of zinc powders, respectively. SEM photograph was taken with scanning electron microscope at the required magnification at room temperature. The working distance of 25 mm was maintained. X-ray Diffractograms were
Fig. 3. The different species of metals in solution were calculated with a chemical equilibrium computer program (MINEQLþ).
obtained in the 2q range of 20e80 using a 0.02 step and acquisition time of 2 s/step. Finally, the size distributions of the same samples were also analyzed by Malvern Laser Particle Size Analyzer. Measurement was delayed until 2 min after sample insertion into the fluid module, to allow the escape of air bubbles formed by sample loading. Grain size distribution was calculated from Fraunhofer or Mie theories, the latter using an optical model depending on the nature of the material. 3. Theory and calculation 3.1. Leaching In this process, zinc and lead were selectively dissolved in sodium hydroxide according to reactions (2)e(3).
ZnOðin the flue dustsÞ þ 2NaOH þ H2 O/Na2 ZnðOHÞ4ðaqÞ
(2)
PbOðin the flue dustsÞ þ 2NaOH þ H2 O/Na2 PbðOHÞ4ðaqÞ
(3)
NaOH could be regenerated during the following electrolysis section. However, large amounts of sodium hydroxide were needed in the leaching section due to the forming of a series of hydroxy-
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complexes. Taking Zn as an example, four different species can be found in the NaOH solution:
Code
Zn2þ þ OH ¼ ½ZnOHþ 2þ
Zn
2þ
Zn
þ 2OH
þ 3OH
(4)
0
¼ ZnðOHÞ2 ¼ ZnðOHÞ3
2 Zn2þ þ 4OH ¼ ZnðOHÞ4
(5) (6) (7)
Given the highest total zinc and lead concentration, the amount of soda hydroxide should exceed 2.7 M according to a rough estimate. In our previous experiments, the solubility of Cu and Al was found to be depressed by the presence of Zn and Pb. However, extremely high levels of copper in these flue dusts made it necessary to determine the optimum pH range for extracting Zn/Pb over Cu. Therefore, a computer simulation was performed using MINEQLþ. Fig. 3 shows that the presence of Zn½OH2 4 as the main zinc species for pH values higher than 14.4. In addition, Pb½OH2 4 became the main lead species progressively as the pH value rose gradually from 14.4 to 14.7, while most of Cu began to dissolve at a higher pH value. Hence, NaOH concentrations of 3 Me5 M (pH ¼ w14.4ew14.7) were applied in this process. 3.2. Electrolysis The electrochemical recoveries of zinc and lead at a cathode occur in the alkaline media, with the following reactions: ZnðOHÞ2 4 þ 2e /Zn þ 4OH
(8)
PbðOHÞ2 4 þ 2e /Pb þ 4OH
(9)
By the Nernst Equation (NaOH ¼ 5 M, Zn ¼ 35 g/L, Pb ¼ 6 g/L): q EZn2þ =Zn ¼ EZn þ 2þ =Zn
RT h 2þ i ln Zn ¼ 1:255 V nF
(10)
EPb2þ =Pb ¼ Eq
RT h 2þ i ln Pb ¼ 0:68 V nF
(11)
Pb
2þ
=Pb
þ
Table 2 Experimental runs and response analysis according to Yates’ algorithm.
1 a b ab c ac bc abc
Variables studied
Zn%
A
B
C
e þ e þ e þ e þ
e e þ þ e e þ þ
e e e e þ þ þ þ
80.47 84.88 83.91 88.52 71.36 79.93 74.02 83.64
Yates’ analysis
Effects
1
2
3
165.35 172.43 151.29 157.66 4.41 4.61 8.57 9.62
337.78 308.95 9.02 18.19 7.08 6.37 0.2 1.05
646.73 27.21 13.45 1.25 28.83 9.17 0.71 0.85
e 6.80 3.36 0.31 7.21 2.29 0.18 0.21
Although it seems that Eq. (11) must take place first, the calculation did not consider the influence of overpotential. Hence, the cyclic voltammograms were measured to complement the above analysis, as shown in Fig. 4. The voltammograms were initiated at 0.5 V/SCE, scanned in the negative direction and reversed at 2.0 V/SCE in the positive direction. The current increased progressively to the maximum value, where it was reversed. Thus, this gave rise to a decrease in current which subsequently reached zero and the current then became anodic corresponding to the dissolution of the deposited zinc and lead. Furthermore, lead was deposited before zinc, as indicated by the appearance of a cathodic peak at about 0.9 to 1.1 V/SCE. The height of the peak increased with Pb concentration in the electrolyte, it became non-appreciable at Pb concentrations of less than 1 g/L. Fig. 4 shows that zinc deposition occurs at a cathodic potential of 1.54 V/SCE at room temperature. Therefore, when lead level exceeds 1 g/L, the selective deposition of lead and zinc is possible considering their difference in redox potentials of about 0.6 V (Suzuki et al., 1995). Since in these circumstances, the crossover potential was equal to the metal/metal-ion reversible potential, it can be assumed that the equilibrium potential values for Zn/Zn2þ and Pb/Pb2þ were around 1.5 V/SCE and 0.9 V/SCE, respectively. These values were more negative than that calculated from Nernst equation (Eqs. (10) and (11)), due to crystallization overpotential related to the substrate.
4. Results and discussion 4.1. Leaching
Fig. 4. Cyclic voltammograms of lead (1e4 g) and zinc (35 g/L) electrowinning in the alkaline leach solution.
Leaching of metals from the dusts is a function of leaching time, concentrations of alkaline solution, phase ratios, and leaching temperature. Firstly, dissolution of Zn and Pb proved to be insignificant below a NaOH concentration of 3 M, then steadily increased with increasing NaOH concentration, attaining a maximum at 4e 5 M. Secondly, the temperature also strongly influences the extraction of zinc and lead. At room temperature, the maximum leaching of Zn and Pb (68e76%) is obtained when the initial concentration of NaOH in the leach solution is 5 M. The leaching efficiencies increased considerably to 80e92% at higher temperatures (>65 C). No significant differences were found over the range of 80 Ce90 C. Hence, a leach temperature range of 65 Ce80 C was selected in order to avoid significant loss of NaOH at higher temperatures. In addition, lower phase ratios of solid to liquid improve the surface contact between the solution and the dusts thus facilitating dissolution of Zn and Pb. However, further decreasing the phase ratios would reduce the Zn and Pb concentrations in the resultant leach solution. Therefore, a phase ratio of 100e125 g/L
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was considered suitable. Furthermore, relatively longer contact time between the dust particles and the liquid was required to obtain maximum extraction of Zn and Pb. One and half hour was found to be sufficient; prolonging the leach times gave only small increase in dissolution. Factorial design and analysis of experiments were used in order to determine the main effects and interactions of the leaching factors (Box et al., 1978). Three quantitative variables were investigated at two levels, as shown in Table 2. These were NaOH concentration (A), leaching temperature (B), and solid/liquid ratio (C). The levels of the variables are given as follows: A [4 (), 5 (þ)] mol/ L, B [65 (), 80 (þ)] C, C [100 (), 125 (þ)] g/L. Reaction time was kept fixed at 1.5 h. The main response under investigation was the percentage of metal (Zn/Pb) recovery in the leach liquor. Table 2 presents that the influence of the parameters on Zn leaching followed the order: S/L ratio > NaOH concentration > leaching temperature > interaction between NaOH concentration and S/L ratio. The other factors, such as interaction between leaching temperature and NaOH concentration, interaction between S/L ratio and leaching temperature, and three-factor interaction have a minor effect on the zinc leaching. Moreover, the same trend in the lead leaching behavior has also been found. In this leaching process, together with Zn and Pb, Al and Cu may also be dissolved in strong alkaline solution. However, the solubility of Cu was found to be negligible in the presence of zinc and lead, and the concentration of Cu in leach solutions is lower than 0.5 g/L. The typical concentration of Al in leach solution was also depressed by the Zn and Pb, but it may be accumulated when the leach solutions are recycled. Hence, part of Al should be removed in certain stage of recycling. Under the optimum condition, the typical contents of the leach solution were (g/L) 35e40 Zn, 6e8 Pb, 0.3e0.8 Al, and 0.2e0.5 Cu. The dissolved copper is removed from the leach liquor by lead cementation at a Pb/Cu molar ratio of 2 within 15 min, and the residual Cu in the solution are 10e50 mg/L which is fully convenient for the following electrolysis of lead and zinc. 4.2. Recovery of Pb by low current density electrolysis According to the analysis and calculation in Section 3, Pb could be recovered over Zn at low current densities (Table 3), which correspond to higher cathodic potential values shown in Fig. 4. Average cell voltages and CE values reported in Table 3 were utilized to calculate the specific energy consumption (PC) by the following equation:
PC ¼
Cell voltage q CE
Fig. 5. SEM micrograph of the produced lead with EDS of selected particles.
The result solution of low current density electrolysis contained Zn ¼ 35e40 g/L, Pb < 1 g/L, Cu ¼ 0.01e0.04 g/L, Al ¼ 0.3e0.8 g/L. Hence, the cementation by adding zinc powders, were carried out to reduce the Pb to low levels (<100 mg/L), as presented in Table 4. The optimum conditions for this cementation were found to be: stoichiometric 1.15 zinc powders (30e50 mm), 35e50 C, and 1.5 h. This Pb level would not damage the purity of zinc powders during the following section, because the cathodic contamination by lead was counteracted by the diffusion limiting current density. Furthermore, during the electrowinning process, diffusion control was rapidly achieved and the Zn2þ electrowinning tended to be predominant.
(12) 4.3. Production of zinc powders by pulse current
where q is electrochemical equivalent of metal, q-Zn ¼ 1.220 g/(Ah), q-Pb ¼ 3.855 g/(Ah). In such way 83e87.5% lead was recovered, and after dissolution lead deposits were analyzed using ICP-OES for purity with respect to the minor elements. Chemical analysis shows the purity of the lead deposits (w97.18%) and the presence of only traces of copper (w1.18%), zinc (w0.72%) and aluminum (w0.87%) elements. Fig. 5 shows the identified elements in the sample by SEM-EDS. The data confirmed the presence of lead as the main element.
Table 3 Electrowinning conditions for lead and zinc recovery.
Pb Pb Zn
5
Cell voltage (V)
Current density (A/m2)
CE (%)
PC (kWh/kg)
1.58 1.90 2.70e3.18
150 250 1000e1500
74 85 80e93
0.55 0.58 2.38e3.26
4.3.1. CE and cell voltage Fig. 6 demonstrates that the average cell voltage increased from w2.9 V at Ton ¼ 5 msew3.2 V at Ton ¼ 15e25 ms, while it decreased from w3.2 V at Toff ¼ 5 ms to w2.6 V at Toff ¼ 20 ms. These behaviors may be related to the decline in the electrolyte electrical conductivity resulting from a longer current-on time.
Table 4 Results of the cementation experiments (1.15 Zn/Pb ratio, 50 C). Time (min)
Removal efficiency of Pb (%) Zn powders (30e50 mm)
Zn powders (60e90 mm)
20 40 60 90
59.05 66.55 82.28 91.83
49.38 52.50 66.64 81.22
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Fig. 6. Cell voltages alkaline zinc electrowinning at different conditions: (a) Toff ¼ 5 ms; (b) Ton ¼ 15 ms.
Table 5 The CE and PC of zinc powder production at different conditions.
Variable (ms) CE (%) PC (kWh/kg)
Toff ¼ 5 ms; Ton ¼ 5e25 m
Ton ¼ 15 ms; Toff ¼ 5e20 m
5 96.2 2.48
5 69.9 3.73
15 69.9 3.73
20 67.2 3.89
25 58.8 4.51
10 81.0 2.93
15 92.86 2.39
20 76.12 2.81
Table 5 presents that sequentially increasing Ton from 5 to 25 ms results in a progressive decrease in the CE down to 58.8%. The explanation for this can be that increasing the Ton does not feed a
sufficient and constant amount of zinc to the cathode, which lowered the reduction rate and consequently the CE. On the contrary, the increase in Toff from 5 to 15 m exerts a positive effect on the CE, and the maximum value (92.86%) was achieved at Toff ¼ 15 ms. Further increase in Toff decreased the CE substantially, this decrease may be attributed to the adsorption of [OH] species on the cathode surface, which blocked the active sites of cathode. The CE and the cell voltage were used to compute the PC, as discussed in the section 4.2. Table 5 shows a negative effect of pulse current, compared with direct current, on the PC. However, this adverse effect could be counteracted by adjusting the Ton and Toff.
Fig. 7. SEM images of zinc powders produced at different conditions: (a) Ton ¼ 5, Toff ¼ 5; (b) Ton ¼ 15, Toff ¼ 5; (c) Ton ¼ 15, Toff ¼ 15; (d) Ton ¼ 15, Toff ¼ 20.
Please cite this article in press as: Qiang, L., et al., Sequential stepwise recovery of selected metals from flue dusts of secondary copper smelting, Journal of Cleaner Production (2014), http://dx.doi.org/10.1016/j.jclepro.2014.03.085
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Fig. 8. Surface morphologies of zinc powders produced at T ¼ 35 C and average current density ¼ 1000 A/m2: (a) direct current electrolysis (b) pulse current electrolysis.
For example, at Ton ¼ 15 and Toff > 10, the PC was 2.4e3.0 kWh/kg, less than the typical industrial value of 3.1e3.2 kWh/kg (Gurmen and Emre, 2003; St-Pierre and Piron, 1990). In addition, the actual rate for zinc electrowinning could be promoted considerably by pulse electrolysis. Another main practical advantage of pulse electrolysis resided in its potential to improve the properties of the product, which is discussed in Section 4.3.2e4.3.3. 4.3.2. Powder characteristic The influence of Ton and Toff on the morphologies of zinc powders, with constant pulse current density, are illustrated in Fig. 7. The zinc powders obtained at Ton ¼ 5 ms were irregular, and relatively rough, with a morphology constituted by a mixture of wide leaf-like particles and boulder deposits. By contrast, dendrites and secondary growth in Zn-Ton ¼ 15 ms were more evidence. It was suggested that the reduction of Zn ions at a higher Ton inhibited the diffusion of adatoms across the surface into the proper sites of the growing crystal lattice. Subsequently, increasing Toff to 15 ms resulted in a rise in grain size of zinc powders. Zinc particles with an average grain size of 43.1 mm were formed at Ton ¼ 15 ms, Toff ¼ 15 ms. This increase in grain size can be explained by a reduced number of nucleation sites caused by the lower overpotential at a longer Toff. However, the zinc powders obtained with further increasing Toff (20 ms) showed a reduced size as indicated in Fig. 7d. This result could be due to the polarization resulting from the [OH] absorbed. Overall, the pulse electrolysis of Ton ¼ 15 ms and Toff ¼ 10 ms gave the best performance in terms of homogeneity and average size of zinc powders (Fig. 8b), and in such way 88e92% Zn was recovered from the leach solution.
4.3.3. Comparison of pulse current and direct electrolysis Electron micrographs of the surface morphology of zinc powders obtained by direct electrolysis and pulse electrolysis are displayed in Figs. 8a and b, respectively. Both of them are dendrite-like and highly porous; the pulse electrolytic zinc powders are more homogeneous and finer. 90% of the direct electrolytic powders are smaller than 235 mm, whereas 90% of the powders produced by pulse current are smaller than 101 mm (Table 6). From XRD it can be seen that the characteristic (101), (100), (002), (102), (103) and
Table 6 Comparisons of pulse current and direct current electrolysis in zinc powder production. Particle Sizea (mm)
D D D D
[3,2] [4,3] [50] [90]
Pulse current
Direct current
Ton ¼ 15 ms; Toff ¼ 10 ms; Average current density ¼ 1000 A/m2
Current density ¼ 1000 A/m2
36.3 60.1 50.9 101
63.3 118.0 92.0 235
a The reported value is typically an equivalent spherical diameter. D[3,2]:The surface diameter; D[4,3]: The volume mean diameter; D[90]: The diameter where 90% of the distribution has a smaller particle size and ten percent has a larger particle size; D[50]: The diameter where 50% of the distribution is above and 50% is below.
Fig. 9. Microstructure analysis of zinc powders produced at T ¼ 35 C and current density ¼ 1000 A/m2: (a) direct current electrolysis (b) pulse current electrolysis.
Please cite this article in press as: Qiang, L., et al., Sequential stepwise recovery of selected metals from flue dusts of secondary copper smelting, Journal of Cleaner Production (2014), http://dx.doi.org/10.1016/j.jclepro.2014.03.085
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(110) peaks of a powder sample from the material shown in Fig. 8a are the same as those of a direct current sample in Fig. 8b (Fig. 9a, b). By comparison, irregular peaks are decreased and crystallization of particles is facilitated by using pulse current. This result could be due to the polarization resulting from the pulse electrolysis, which provides sufficient energy for the crystallization. 4.4. Removal of Al and Cl The solution resulting from pulse electrolysis contains Al (0.3e 0.8 g/L) and Cl (4.5e7 g/L). They may be accumulated when the leach solutions are recycled. Maximum allowable concentrations of Al and Cl were found to be 7 g/L and 25 g/L in this process, respectively. Further increasing the amount of Al and Cl could affect adversely the process. At higher levels (>7 g/L), Al could decrease the electrolyte conductivity, thus inhibiting the zinc ion reduction, and consequently increases the cell voltage and PC. Thus, Al was separated from the electrolyte by adding CaO due to the following reaction:
3CaO þ 2AlðOHÞ 4 þ 3H2 O ¼ Ca3 AlðOHÞ6 2 þ 2OH :
(13)
When sufficient CaO was added to make the molar ratios of Ca to Al in the solution equal to 1.8e2.0, 65e72% Al could be recovered from the solution within 4 h. Subsequently, an evaporation step was carried out at 90 C. In this step, NaOH concentration was increased up to 450 g/L allowing the NaCl precipitation, and consequently 92e93% Cl was removed from the electrolyte. The concentrated electrolyte was recycled to the leaching. 5. Conclusion A new approach was suggested to treat flue dusts formed during secondary copper smelting, with the aim of maximizing metals recovery and minimizing generated residues. The method developed for processing the flue dust comprises several steps: Leaching of dusts using a 5 M NaOH solution at 100 g/L S/L ratio and at 80 C for about 1.5 h. These conditions allow a high recovery of Zn and Pb, while minimizing the co-extraction of Cu. Electrolysis of leach liquor at current densities of 100e250 A/m2 which allows the recovery of 83e87.5% Pb from the solution. Pulse current electrolysis of the result solution at Ton ¼ 15 ms and Toff ¼ 10 ms to obtain homogeneous zinc powders, and a recovery of 88e92% of Zn. Precipitation of Cl and Al when their concentrations exceeded 25 g/L and 7 g/L, respectively, which avoids the accumulation of Al and Cl when reusing the solution in the alkaline leaching stage. Integration of alkaline media leaching and two-stage electrolysis can make the whole process economically attractive due to the value-added by-products and environmental friendly due to the nearly-closed cycle. In particular, pulse current makes the zinc powder more uniform and finer, thus improving its marketability. Furthermore, this product could be an attractive material for organic synthesis, brass manufacture and the battery industry. This integration also allows a high level Cl in the electrolyte, simplifying the procedure. The proposed process might contribute to a more sustainable and cost-effective management for other industrial wastes such as hot galvanizing slag, spent catalyst and automobile shredder scrap, but these areas still deserve further attention. Moreover, future research is also needed that explores ways to reduce NaOH consumption and leaching temperature.
Acknowledgments This work was financially supported by China Ministry of Environmental Protection (No. 2013467025), and by National Natural Science Funds (No. 51278350). Additionally, authors are thankful to the Department of Environmental Science and Technology, Tongji University for equipment support. Authors would also like to appreciate kind cooperation of Helena M.V.M. Soares in Porto University for extending necessary help. References Box, G.E.P., Hunter, W.G., Hunter, J.S., 1978. Statistics for Experiments. John Willey, New York. Chandrasekar, M.S., Shanmugasigamani, Pushpavanam, Malathy, 2010. Synergetic effects of pulse constraints and additives in electrodeposition of nanocrystalline zinc: corrosion, structural and textural characterization. Mater. Chem. Phys. 124, 516e528. Dutra, A.J.B., Paiva, P.R.P., Tavares, L.M., 2006. Alkaline leaching of zinc from electric arc furnace steel dust. Min. Eng. 19, 478e485. Dvorak, P., Jandova, J., 2006. Zinc recovery from chloride-bearing galvanizing waste. Acta. Metall. Slovaca 12, 90e94. Frenay, J., Ferlay, S., Hissel, J., 1986. Zinc and lead recovery from EAF dusts by caustic soda processElectric Furnace Proceedings, Treatment Options for Carbon Steel Electric Arc Furnace Dust. Iron Steel Soc. 43, 417e421. Ghavami, R.K., Rafiei, Z., 2006. Performance improvements of alkaline batteries by studying the effects of different kinds of surfactant and different derivatives of benzene on the electrochemical properties of electrolytic zinc. J. Pow. Sour. 162, 893e899. Gupta, M.K., Gupta, B.L., Raghavan, R., 1989. Investigation of use of zinc dross as a substitute for zinc dust in solution purification in zinc electrowinning hydrometallurgical plant. Hydrometallurgy 22, 379e391. Guresin, N., Topkaya, Y.A., 1998. Dechlorination of a zinc dross. Hydrometallurgy 49, 179e187. Gurmen, S., Emre, M., 2003. A laboratory-scale investigation of alkaline zinc electrowinning. Min. Eng. 16, 559e562. Ju, S.H., Zhang, Y.F., Zhang, Y., Xue, P.Y., Wang, Y.H., 2011. Clean hydrometallurgical route to recover zinc, silver, lead, copper, cadmium and iron from hazardous jarosite residues produced during zinc hydrometallurgy. J. Hazard. Mater. 192, 554e558. Langova, S., Riplova, J., Vallova, S., 2007. Atmospheric leaching of steel-making wastes and the precipitation of goethite from the ferric sulphate solution. Hydrometallurgy 87, 157e162. Li, J.S., Zhu, J.T., Yang, H.B., Yang, S.F., 2011. Status on the treatment of EAF dust in China. Henan Metall. 19, 1e5. Martell, A.E., Smith, R.M., 2004. NIST Standard Reference Database 46 Version 8.0. NIST Critically Selected Stability Constants of Metal Complexes Database. US Department of Commerce, National Institute of Standards and Technology. Oustadakis, P., Tsakiridis, P.E., Katsiapi, A., Agatzini-Leonardou, S., 2010. Hydrometallurgical process for zinc recovery from electric arc furnace dust (EAFD). Part I: characterization and leaching by diluted sulphuric acid. J. Hazard. Mater. 179, 1e7. Pinto, I.S.S., Soares, H.M.V.M., 2012. Selective leaching of molybdenum from spent hydrodesulphurization catalysts using ultrasound and microwave methods. Hydrometallurgy 129e130, 19e25. Pinto, I.S.S., Soares, H.M.V.M., 2013. Recovery of molybdates from an alkaline leachate of spent hydrodesulphurisation catalysteproposal of a nearly-closed process. J. Clean. Prod. 52, 481e487. Saber, Kh., Koch, C.C., Fedkiw, P.S., 2003. Pulse current electrodeposition of nanocrystalline zinc. Mater. Sci. Eng. A 341, 174e181. Schecher, W.D., McAvoy, D.C., 2003. MINEQLt A Chemical Equilibrium Modeling System, Version 4.5 for Windows. User’s Manual. Hallowell, Maine. St-Pierre, J., Piron, D.L., 1990. Electrowinning of zinc from alkaline solutions at high current densities. J. Appl. Electrochem 20, 163e165. Suzuki, R., Li, W.H., Schwarz, M., Nobe, K., 1995. Segmented porous-electrode flow reactors for the electrochemical treatment of commingled metal plating wastes. Plat. Surf. Finish 82, 58e65. 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. J. Hazard. Mater. 192, 1100e1107. Tsakiridis, P.E., Oustadakis, P., Katsiapi, A., Agatzini-Leonardou, S., 2010. Hydrometallurgical process for zinc recovery from electric arc furnace dust (EAFD). Part II: downstream processing and zinc recovery by electrowinning. J. Hazard. Mater. 179, 8e14. Youssef, K.M., Koch, C.C., Fedkiw, P.S., 2008. Influence of pulse plating parameters on the synthesis and preferred orientation of nanocrystalline zinc from zinc sulfate electrolytes. Electrochim. Acta 54, 677e683. Zhao, Y.C., Stanforth, R., 2001. Selective separation of lead from alkaline zinc solution by sulfide precipitation. Sep. Sci. Technol. 36, 2561e2570.
Please cite this article in press as: Qiang, L., et al., Sequential stepwise recovery of selected metals from flue dusts of secondary copper smelting, Journal of Cleaner Production (2014), http://dx.doi.org/10.1016/j.jclepro.2014.03.085
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Journal of Cleaner Production journal homepage: www.elsevier.com/locate/jclepro
Sequential stepwise recovery of selected metals from flue dusts of secondary copper smelting Li Qiang a, Isabel S.S. Pinto b, Zhao Youcai a, * a b
The State Key Laboratory of Pollution Control and Resource Reuse, Tongji University, Shanghai 200092, China REQUIMTE, Faculty of Engineering, University of Porto, Rua Dr.Roberto Frias, 4200-465 Porto, Portugal
a r t i c l e i n f o
a b s t r a c t
Article history: Received 4 August 2013 Received in revised form 22 March 2014 Accepted 24 March 2014 Available online xxx
A novel and clean process was developed to convert flue dusts of secondary copper smelters into valueadded products. The process consists of three main steps: NaOH leaching, two-stage combined electrolysis, and deep purification of recycle solution. Initially, 80e92% Zn and Pb were dissolved in 5 M NaOH at 80 C, whereas Cu was concentrated in the residue. Yates’ algorithm was used to determine the main effects and interactions of the leaching factors. The leach solution was then electrolyzed at 100e250 A/m2 and high purity Pb (>97%) was separated from the alkaline solution. Subsequently, a pulsed current was introduced to obtain ultrafine zinc powders (w30 mm) with the best performance occurred at Ton (current-on) ¼ 15 ms, and Toff (current-off) ¼ 10 ms. Finally, Cl and Al were precipitated by evaporation-condensation and CaO addition, respectively. The combination of alkaline media and pulse current enables the green route without distilled water washing and surfactant addition, thus minimizing effluent emissions. This environmentally friendly method has promise to treat other industrial wastes due to its low cost and simple design. Ó 2014 Elsevier Ltd. All rights reserved.
Keywords: Green process Metal recovery Yates’ algorithm Ultrafine zinc powder High pure lead Hazardous wastes
1. Introduction Metallurgical industries generate vast quantities of different types of wastes such as electric arc furnace (EAF) dust, basic oxygen furnace (BOF) sludge, jarosite residue and flue dust of secondary copper smelting (Dvorak and Jandova, 2006; Ju et al., 2011). On the basis of rough statistics, no less than one million tons of these wastes have been produced in China per year (Li et al., 2011). These wastes could be used to recover metallic values or they may be disposed of. However, the disposal of such material is now becoming expensive due to increasingly stringent environmental regulations. Furthermore, the chemical nature of these dust/ash particles is such that these are classified as hazardous waste under the US Environmental Protection Agency classification. In view of the above, there has been an increasing interest in developing processes for the recovery of metals from these wastes. Usually, pyrometallurgical and hydrometallurgical processes are employed for treating such residues. The pyrometallurgical processes face the problems of generation of worthless residues and
* Corresponding author. Fax: þ86 21 65982684. E-mail addresses: [email protected], [email protected] ( Zhao Youcai).
costly equipment investment as well as quite high energy consumption (Pinto and Soares, 2012, 2013). Thus, much more attention has been paid to hydrometallurgical processes. For example, Dutra et al. (2006) extracted the zinc present in EAF dust with different alkaline leaching techniques. Oustadakis and coauthors developed a process of H2SO4 leaching, solvent extraction and electrowinning to recover zinc from EAF dust (Oustadakis et al., 2010; Tsakiridis et al., 2010). Langova et al. (2007) examined the sulfuric acid leaching of EAF dust and studied the influence of acid concentration, temperature, time, and liquid/solid ratio. Zinc extraction reached almost 100% and iron extraction exceeded 90% in 3 M H2SO4 at 80 C and liquid/solid ratio of 5 after 6 h. Furthermore, they found that a good selectivity with regard to zinc was achieved with 0.1e0.3 M H2SO4 at 80 C. Ju et al. (2011) proposed a clean hydrometallurgical route to recover zinc, silver, lead, copper, cadmium and iron from hazardous jarosite residues produced by zinc hydrometallurgy, whereas Trung et al. (2011) concentrated on the zinc leaching from fine-grain BOF sludge in acid medium. However, there is still not enough information about how to process wastes from the secondary Cu industry. The flue dusts used in this work were obtained from the city of Fuyang, which is situated in Zhejiang, a developed southeastern province in China; it is one of the biggest national centers for secondary Cu industries. The composition of the dusts varies
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Please cite this article in press as: Qiang, L., et al., Sequential stepwise recovery of selected metals from flue dusts of secondary copper smelting, Journal of Cleaner Production (2014), http://dx.doi.org/10.1016/j.jclepro.2014.03.085
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L. Qiang et al. / Journal of Cleaner Production xxx (2014) 1e8
considerably, and it is dependent not only on the waste used, but also on the operating conditions. However, some general trends have been noted. For example, most of these dusts are rich in Zn, Cu, Cl, Pb, and Al. It is advisable to leach the dusts with caustic soda, because the chlorides will not be tolerated in the acid electrolyte, as even a very small amount can cause severe corrosion problems and thereby damage the electrolysis (Gupta et al., 1989; Guresin and Topkaya, 1998). Moreover, the washing section for removing Cl produces secondary pollutants and complicates the process. During caustic soda process, however, the chloride concentration increased over 25 g/L without any trouble being observed (Frenay et al., 1986). In particular no chlorine evolution was noted in the electrolysis cells, and formation of hypochlorite would be expected from the following reaction:
Cl þ H2 O/HClO þ Hþ þ 2e :
(1)
In addition, most of copper (>90%) is left in the residue when using this alkaline leach reagent. Furthermore, in the alkaline zinc electrowinning process hydrogen evolution on the cathode surface is even more impeded than the one in acidic electrowinning. Current efficiencies (CE) of 97e99% can thus be reached when high current densities are applied (Gurmen and Emre, 2003; St-Pierre and Piron, 1990). This behavior promotes the generation of more profitable products with a lower energy cost. On the basis of the above analysis, a hydrometallurgical route is presented in this work, as shown in Fig. 1, for the metal recovery from the flue dusts of secondary copper smelters. In the alkaline zinc electrowinning, Na2S has been used to remove Pb from the leach solution, and this approach can recover lead selectively and quantitatively (Zhao and Stanforth, 2001). However, it is time consuming to separate Pb from alkaline solution, and Na2S is a toxic agent. Therefore, the present research focuses on the lead recovery from the alkaline media by low current densities electrolysis. In the conventional process, surfactants are widely applied to control the
size and morphology of zinc powders (Ghavami and Rafiei, 2006); however the accumulation of these additives could damage the recycle solution quality and increase effluent emissions. Hence, other potential methods are clearly needed to produce ultrafine zinc powders with uniform size. In this work, pulse electrolysis is employed to obtain zinc powder due to its variety of mass transport situations and electrocrystallization conditions. However, the majority of studies on pulse electrolysis have focused on its potential advantages in the electroplating industry (Chandrasekar et al., 2010; Saber et al., 2003; Youssef et al., 2008). Information is lacking regarding its use in powder production. This work describes a novel and efficient method to extract Pb from an alkaline leaching solution of the sample. Attempt was also made to produce uniform and fine zinc powders with pulse current. Subsequently, we reported the techniques for removing Cl and Al from the resultant alkaline electrolyte. 2. Material and methods 2.1. Characterization of the flue dusts The sample was dried and homogenized before the leaching experiments, and its chemical composition is listed in Table 1. The chemical analysis was determined by ICP-OES (AGILENT 720ES) and Ion Chromatography (ICS-1000). It demonstrates that the major elements present in the sample are zinc, copper, lead, aluminum and chlorine. An XRD (BRUKER D8 ADVANCE) analysis of the asreceived dust is displayed in Fig. 2. It shows that most of metal elements are in combination with oxygen, but there can also be chloride. 2.2. Computer chemical simulations Chemical speciation calculations were carried out using the MINEQLþ software (version 4.5) (Schecher and McAvoy, 2003), and the species distribution of metals at different pH values were determined. Metal speciation analysis with MINEQLþ generates chemical equilibrium concentrations of all species being considered in the model by the program reactions, based on component stability constants (Martell and Smith, 2004) and total molar metal concentrations. For computer simulations, total molar metal concentrations were calculated considering a solid to liquid (S/L) ratio of 100 g/L and assuming that the total amount of metals in the dust were in solution. The simulations allow predicting the optimum pH range to solubilize the target metals. 2.3. Procedure The proposed process consists of three main stages, including NaOH leaching, two-stage combined electrolysis, and deep purification of recycle solution. With this method, the stepwise extraction of Cu, Pb, Zn and Al was achieved. Moreover, ICP-OES, XRD, SEM-EDS (PHILIPS XL30), IC, particle size analyses (Malvern 3000), and electrochemical methods were used in this study. 2.3.1. Leaching The leaching processes were carried out at 30e90 C under atmospheric pressure. Specific amounts of flue dust samples were Table 1 Chemical analysis of the main elements present in the dust.
Fig. 1. General scheme for treating flue dusts from secondary copper smelters in this process.
Element
Zn
Cu
Pb
Al
Cl
Weight (%)
40.21
7.53
6.62
2.58
8.47
Please cite this article in press as: Qiang, L., et al., Sequential stepwise recovery of selected metals from flue dusts of secondary copper smelting, Journal of Cleaner Production (2014), http://dx.doi.org/10.1016/j.jclepro.2014.03.085
L. Qiang et al. / Journal of Cleaner Production xxx (2014) 1e8
3
Fig. 2. Mineralogical phases of flue dusts used.
added into a glass reactor containing a given volume of 3e6 M NaOH solution, and then the reactor rotated with an orbital shaker for 1e2 h (15 rpm). A series of leach tests were run, and the effects of variables on zinc recovery were assessed using Yates’ algorithm. Finally, the supernatant and the residues were analyzed by ICP-OES. 2.3.2. Electrolysis In electrowinning, stainless steel electrodes were used as both anode and cathode. Moreover, the anode to cathode distance was kept at 3 cm and all electrowinning experiments were carried out at room temperature (30e50 C). Current, cell voltage, and metal concentrations were recorded during the electrolysis, while at least two replicates were used for every measurement. Finally, the cathode CE was calculated from the mass gain obtained after electrolysis. Electrolysis of lead was attained by applying a low current density of (100e250 A/m2). At the end of electrowinning, Pb concentration in the final solution was quantified by ICP-OES to determine the efficiency of metal recovery. The purity of the product was also analyzed by ICP-OES and SEM-EDS. Zinc powders were produced by pulse direct current. The current-on time (Ton) and the current-off time (Toff) were optimized to obtain fine and homogeneous particles. The optimization was performed through a series of experiments in which Ton was varied from 5 to 25 ms and Toff from 5 to 20 ms at a constant pulse current density (1500e1800 A/m2). 2.4. Analysis method 2.4.1. Electrochemical measurements A CHI600D electrochemical workstation was used for the polarization studies, and all the polarization measurements were performed in a three-electrode Plexiglas cell. A couple of stainless steel plates were employed as working and auxiliary electrodes, while reference electrode was a saturated calomel electrode (SCE) with an appropriate lugging capillary. All potentials were recorded with respect to SCE. Cyclic voltammetries were carried out by initiating scans at 0.5 V versus SCE at a rate of 5 mV s1. 2.4.2. Powder examination SEM and XRD were used to characterize the surface morphology and the preferred crystal orientation of zinc powders, respectively. SEM photograph was taken with scanning electron microscope at the required magnification at room temperature. The working distance of 25 mm was maintained. X-ray Diffractograms were
Fig. 3. The different species of metals in solution were calculated with a chemical equilibrium computer program (MINEQLþ).
obtained in the 2q range of 20e80 using a 0.02 step and acquisition time of 2 s/step. Finally, the size distributions of the same samples were also analyzed by Malvern Laser Particle Size Analyzer. Measurement was delayed until 2 min after sample insertion into the fluid module, to allow the escape of air bubbles formed by sample loading. Grain size distribution was calculated from Fraunhofer or Mie theories, the latter using an optical model depending on the nature of the material. 3. Theory and calculation 3.1. Leaching In this process, zinc and lead were selectively dissolved in sodium hydroxide according to reactions (2)e(3).
ZnOðin the flue dustsÞ þ 2NaOH þ H2 O/Na2 ZnðOHÞ4ðaqÞ
(2)
PbOðin the flue dustsÞ þ 2NaOH þ H2 O/Na2 PbðOHÞ4ðaqÞ
(3)
NaOH could be regenerated during the following electrolysis section. However, large amounts of sodium hydroxide were needed in the leaching section due to the forming of a series of hydroxy-
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complexes. Taking Zn as an example, four different species can be found in the NaOH solution:
Code
Zn2þ þ OH ¼ ½ZnOHþ 2þ
Zn
2þ
Zn
þ 2OH
þ 3OH
(4)
0
¼ ZnðOHÞ2 ¼ ZnðOHÞ3
2 Zn2þ þ 4OH ¼ ZnðOHÞ4
(5) (6) (7)
Given the highest total zinc and lead concentration, the amount of soda hydroxide should exceed 2.7 M according to a rough estimate. In our previous experiments, the solubility of Cu and Al was found to be depressed by the presence of Zn and Pb. However, extremely high levels of copper in these flue dusts made it necessary to determine the optimum pH range for extracting Zn/Pb over Cu. Therefore, a computer simulation was performed using MINEQLþ. Fig. 3 shows that the presence of Zn½OH2 4 as the main zinc species for pH values higher than 14.4. In addition, Pb½OH2 4 became the main lead species progressively as the pH value rose gradually from 14.4 to 14.7, while most of Cu began to dissolve at a higher pH value. Hence, NaOH concentrations of 3 Me5 M (pH ¼ w14.4ew14.7) were applied in this process. 3.2. Electrolysis The electrochemical recoveries of zinc and lead at a cathode occur in the alkaline media, with the following reactions: ZnðOHÞ2 4 þ 2e /Zn þ 4OH
(8)
PbðOHÞ2 4 þ 2e /Pb þ 4OH
(9)
By the Nernst Equation (NaOH ¼ 5 M, Zn ¼ 35 g/L, Pb ¼ 6 g/L): q EZn2þ =Zn ¼ EZn þ 2þ =Zn
RT h 2þ i ln Zn ¼ 1:255 V nF
(10)
EPb2þ =Pb ¼ Eq
RT h 2þ i ln Pb ¼ 0:68 V nF
(11)
Pb
2þ
=Pb
þ
Table 2 Experimental runs and response analysis according to Yates’ algorithm.
1 a b ab c ac bc abc
Variables studied
Zn%
A
B
C
e þ e þ e þ e þ
e e þ þ e e þ þ
e e e e þ þ þ þ
80.47 84.88 83.91 88.52 71.36 79.93 74.02 83.64
Yates’ analysis
Effects
1
2
3
165.35 172.43 151.29 157.66 4.41 4.61 8.57 9.62
337.78 308.95 9.02 18.19 7.08 6.37 0.2 1.05
646.73 27.21 13.45 1.25 28.83 9.17 0.71 0.85
e 6.80 3.36 0.31 7.21 2.29 0.18 0.21
Although it seems that Eq. (11) must take place first, the calculation did not consider the influence of overpotential. Hence, the cyclic voltammograms were measured to complement the above analysis, as shown in Fig. 4. The voltammograms were initiated at 0.5 V/SCE, scanned in the negative direction and reversed at 2.0 V/SCE in the positive direction. The current increased progressively to the maximum value, where it was reversed. Thus, this gave rise to a decrease in current which subsequently reached zero and the current then became anodic corresponding to the dissolution of the deposited zinc and lead. Furthermore, lead was deposited before zinc, as indicated by the appearance of a cathodic peak at about 0.9 to 1.1 V/SCE. The height of the peak increased with Pb concentration in the electrolyte, it became non-appreciable at Pb concentrations of less than 1 g/L. Fig. 4 shows that zinc deposition occurs at a cathodic potential of 1.54 V/SCE at room temperature. Therefore, when lead level exceeds 1 g/L, the selective deposition of lead and zinc is possible considering their difference in redox potentials of about 0.6 V (Suzuki et al., 1995). Since in these circumstances, the crossover potential was equal to the metal/metal-ion reversible potential, it can be assumed that the equilibrium potential values for Zn/Zn2þ and Pb/Pb2þ were around 1.5 V/SCE and 0.9 V/SCE, respectively. These values were more negative than that calculated from Nernst equation (Eqs. (10) and (11)), due to crystallization overpotential related to the substrate.
4. Results and discussion 4.1. Leaching
Fig. 4. Cyclic voltammograms of lead (1e4 g) and zinc (35 g/L) electrowinning in the alkaline leach solution.
Leaching of metals from the dusts is a function of leaching time, concentrations of alkaline solution, phase ratios, and leaching temperature. Firstly, dissolution of Zn and Pb proved to be insignificant below a NaOH concentration of 3 M, then steadily increased with increasing NaOH concentration, attaining a maximum at 4e 5 M. Secondly, the temperature also strongly influences the extraction of zinc and lead. At room temperature, the maximum leaching of Zn and Pb (68e76%) is obtained when the initial concentration of NaOH in the leach solution is 5 M. The leaching efficiencies increased considerably to 80e92% at higher temperatures (>65 C). No significant differences were found over the range of 80 Ce90 C. Hence, a leach temperature range of 65 Ce80 C was selected in order to avoid significant loss of NaOH at higher temperatures. In addition, lower phase ratios of solid to liquid improve the surface contact between the solution and the dusts thus facilitating dissolution of Zn and Pb. However, further decreasing the phase ratios would reduce the Zn and Pb concentrations in the resultant leach solution. Therefore, a phase ratio of 100e125 g/L
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L. Qiang et al. / Journal of Cleaner Production xxx (2014) 1e8
was considered suitable. Furthermore, relatively longer contact time between the dust particles and the liquid was required to obtain maximum extraction of Zn and Pb. One and half hour was found to be sufficient; prolonging the leach times gave only small increase in dissolution. Factorial design and analysis of experiments were used in order to determine the main effects and interactions of the leaching factors (Box et al., 1978). Three quantitative variables were investigated at two levels, as shown in Table 2. These were NaOH concentration (A), leaching temperature (B), and solid/liquid ratio (C). The levels of the variables are given as follows: A [4 (), 5 (þ)] mol/ L, B [65 (), 80 (þ)] C, C [100 (), 125 (þ)] g/L. Reaction time was kept fixed at 1.5 h. The main response under investigation was the percentage of metal (Zn/Pb) recovery in the leach liquor. Table 2 presents that the influence of the parameters on Zn leaching followed the order: S/L ratio > NaOH concentration > leaching temperature > interaction between NaOH concentration and S/L ratio. The other factors, such as interaction between leaching temperature and NaOH concentration, interaction between S/L ratio and leaching temperature, and three-factor interaction have a minor effect on the zinc leaching. Moreover, the same trend in the lead leaching behavior has also been found. In this leaching process, together with Zn and Pb, Al and Cu may also be dissolved in strong alkaline solution. However, the solubility of Cu was found to be negligible in the presence of zinc and lead, and the concentration of Cu in leach solutions is lower than 0.5 g/L. The typical concentration of Al in leach solution was also depressed by the Zn and Pb, but it may be accumulated when the leach solutions are recycled. Hence, part of Al should be removed in certain stage of recycling. Under the optimum condition, the typical contents of the leach solution were (g/L) 35e40 Zn, 6e8 Pb, 0.3e0.8 Al, and 0.2e0.5 Cu. The dissolved copper is removed from the leach liquor by lead cementation at a Pb/Cu molar ratio of 2 within 15 min, and the residual Cu in the solution are 10e50 mg/L which is fully convenient for the following electrolysis of lead and zinc. 4.2. Recovery of Pb by low current density electrolysis According to the analysis and calculation in Section 3, Pb could be recovered over Zn at low current densities (Table 3), which correspond to higher cathodic potential values shown in Fig. 4. Average cell voltages and CE values reported in Table 3 were utilized to calculate the specific energy consumption (PC) by the following equation:
PC ¼
Cell voltage q CE
Fig. 5. SEM micrograph of the produced lead with EDS of selected particles.
The result solution of low current density electrolysis contained Zn ¼ 35e40 g/L, Pb < 1 g/L, Cu ¼ 0.01e0.04 g/L, Al ¼ 0.3e0.8 g/L. Hence, the cementation by adding zinc powders, were carried out to reduce the Pb to low levels (<100 mg/L), as presented in Table 4. The optimum conditions for this cementation were found to be: stoichiometric 1.15 zinc powders (30e50 mm), 35e50 C, and 1.5 h. This Pb level would not damage the purity of zinc powders during the following section, because the cathodic contamination by lead was counteracted by the diffusion limiting current density. Furthermore, during the electrowinning process, diffusion control was rapidly achieved and the Zn2þ electrowinning tended to be predominant.
(12) 4.3. Production of zinc powders by pulse current
where q is electrochemical equivalent of metal, q-Zn ¼ 1.220 g/(Ah), q-Pb ¼ 3.855 g/(Ah). In such way 83e87.5% lead was recovered, and after dissolution lead deposits were analyzed using ICP-OES for purity with respect to the minor elements. Chemical analysis shows the purity of the lead deposits (w97.18%) and the presence of only traces of copper (w1.18%), zinc (w0.72%) and aluminum (w0.87%) elements. Fig. 5 shows the identified elements in the sample by SEM-EDS. The data confirmed the presence of lead as the main element.
Table 3 Electrowinning conditions for lead and zinc recovery.
Pb Pb Zn
5
Cell voltage (V)
Current density (A/m2)
CE (%)
PC (kWh/kg)
1.58 1.90 2.70e3.18
150 250 1000e1500
74 85 80e93
0.55 0.58 2.38e3.26
4.3.1. CE and cell voltage Fig. 6 demonstrates that the average cell voltage increased from w2.9 V at Ton ¼ 5 msew3.2 V at Ton ¼ 15e25 ms, while it decreased from w3.2 V at Toff ¼ 5 ms to w2.6 V at Toff ¼ 20 ms. These behaviors may be related to the decline in the electrolyte electrical conductivity resulting from a longer current-on time.
Table 4 Results of the cementation experiments (1.15 Zn/Pb ratio, 50 C). Time (min)
Removal efficiency of Pb (%) Zn powders (30e50 mm)
Zn powders (60e90 mm)
20 40 60 90
59.05 66.55 82.28 91.83
49.38 52.50 66.64 81.22
Please cite this article in press as: Qiang, L., et al., Sequential stepwise recovery of selected metals from flue dusts of secondary copper smelting, Journal of Cleaner Production (2014), http://dx.doi.org/10.1016/j.jclepro.2014.03.085
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Fig. 6. Cell voltages alkaline zinc electrowinning at different conditions: (a) Toff ¼ 5 ms; (b) Ton ¼ 15 ms.
Table 5 The CE and PC of zinc powder production at different conditions.
Variable (ms) CE (%) PC (kWh/kg)
Toff ¼ 5 ms; Ton ¼ 5e25 m
Ton ¼ 15 ms; Toff ¼ 5e20 m
5 96.2 2.48
5 69.9 3.73
15 69.9 3.73
20 67.2 3.89
25 58.8 4.51
10 81.0 2.93
15 92.86 2.39
20 76.12 2.81
Table 5 presents that sequentially increasing Ton from 5 to 25 ms results in a progressive decrease in the CE down to 58.8%. The explanation for this can be that increasing the Ton does not feed a
sufficient and constant amount of zinc to the cathode, which lowered the reduction rate and consequently the CE. On the contrary, the increase in Toff from 5 to 15 m exerts a positive effect on the CE, and the maximum value (92.86%) was achieved at Toff ¼ 15 ms. Further increase in Toff decreased the CE substantially, this decrease may be attributed to the adsorption of [OH] species on the cathode surface, which blocked the active sites of cathode. The CE and the cell voltage were used to compute the PC, as discussed in the section 4.2. Table 5 shows a negative effect of pulse current, compared with direct current, on the PC. However, this adverse effect could be counteracted by adjusting the Ton and Toff.
Fig. 7. SEM images of zinc powders produced at different conditions: (a) Ton ¼ 5, Toff ¼ 5; (b) Ton ¼ 15, Toff ¼ 5; (c) Ton ¼ 15, Toff ¼ 15; (d) Ton ¼ 15, Toff ¼ 20.
Please cite this article in press as: Qiang, L., et al., Sequential stepwise recovery of selected metals from flue dusts of secondary copper smelting, Journal of Cleaner Production (2014), http://dx.doi.org/10.1016/j.jclepro.2014.03.085
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Fig. 8. Surface morphologies of zinc powders produced at T ¼ 35 C and average current density ¼ 1000 A/m2: (a) direct current electrolysis (b) pulse current electrolysis.
For example, at Ton ¼ 15 and Toff > 10, the PC was 2.4e3.0 kWh/kg, less than the typical industrial value of 3.1e3.2 kWh/kg (Gurmen and Emre, 2003; St-Pierre and Piron, 1990). In addition, the actual rate for zinc electrowinning could be promoted considerably by pulse electrolysis. Another main practical advantage of pulse electrolysis resided in its potential to improve the properties of the product, which is discussed in Section 4.3.2e4.3.3. 4.3.2. Powder characteristic The influence of Ton and Toff on the morphologies of zinc powders, with constant pulse current density, are illustrated in Fig. 7. The zinc powders obtained at Ton ¼ 5 ms were irregular, and relatively rough, with a morphology constituted by a mixture of wide leaf-like particles and boulder deposits. By contrast, dendrites and secondary growth in Zn-Ton ¼ 15 ms were more evidence. It was suggested that the reduction of Zn ions at a higher Ton inhibited the diffusion of adatoms across the surface into the proper sites of the growing crystal lattice. Subsequently, increasing Toff to 15 ms resulted in a rise in grain size of zinc powders. Zinc particles with an average grain size of 43.1 mm were formed at Ton ¼ 15 ms, Toff ¼ 15 ms. This increase in grain size can be explained by a reduced number of nucleation sites caused by the lower overpotential at a longer Toff. However, the zinc powders obtained with further increasing Toff (20 ms) showed a reduced size as indicated in Fig. 7d. This result could be due to the polarization resulting from the [OH] absorbed. Overall, the pulse electrolysis of Ton ¼ 15 ms and Toff ¼ 10 ms gave the best performance in terms of homogeneity and average size of zinc powders (Fig. 8b), and in such way 88e92% Zn was recovered from the leach solution.
4.3.3. Comparison of pulse current and direct electrolysis Electron micrographs of the surface morphology of zinc powders obtained by direct electrolysis and pulse electrolysis are displayed in Figs. 8a and b, respectively. Both of them are dendrite-like and highly porous; the pulse electrolytic zinc powders are more homogeneous and finer. 90% of the direct electrolytic powders are smaller than 235 mm, whereas 90% of the powders produced by pulse current are smaller than 101 mm (Table 6). From XRD it can be seen that the characteristic (101), (100), (002), (102), (103) and
Table 6 Comparisons of pulse current and direct current electrolysis in zinc powder production. Particle Sizea (mm)
D D D D
[3,2] [4,3] [50] [90]
Pulse current
Direct current
Ton ¼ 15 ms; Toff ¼ 10 ms; Average current density ¼ 1000 A/m2
Current density ¼ 1000 A/m2
36.3 60.1 50.9 101
63.3 118.0 92.0 235
a The reported value is typically an equivalent spherical diameter. D[3,2]:The surface diameter; D[4,3]: The volume mean diameter; D[90]: The diameter where 90% of the distribution has a smaller particle size and ten percent has a larger particle size; D[50]: The diameter where 50% of the distribution is above and 50% is below.
Fig. 9. Microstructure analysis of zinc powders produced at T ¼ 35 C and current density ¼ 1000 A/m2: (a) direct current electrolysis (b) pulse current electrolysis.
Please cite this article in press as: Qiang, L., et al., Sequential stepwise recovery of selected metals from flue dusts of secondary copper smelting, Journal of Cleaner Production (2014), http://dx.doi.org/10.1016/j.jclepro.2014.03.085
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(110) peaks of a powder sample from the material shown in Fig. 8a are the same as those of a direct current sample in Fig. 8b (Fig. 9a, b). By comparison, irregular peaks are decreased and crystallization of particles is facilitated by using pulse current. This result could be due to the polarization resulting from the pulse electrolysis, which provides sufficient energy for the crystallization. 4.4. Removal of Al and Cl The solution resulting from pulse electrolysis contains Al (0.3e 0.8 g/L) and Cl (4.5e7 g/L). They may be accumulated when the leach solutions are recycled. Maximum allowable concentrations of Al and Cl were found to be 7 g/L and 25 g/L in this process, respectively. Further increasing the amount of Al and Cl could affect adversely the process. At higher levels (>7 g/L), Al could decrease the electrolyte conductivity, thus inhibiting the zinc ion reduction, and consequently increases the cell voltage and PC. Thus, Al was separated from the electrolyte by adding CaO due to the following reaction:
3CaO þ 2AlðOHÞ 4 þ 3H2 O ¼ Ca3 AlðOHÞ6 2 þ 2OH :
(13)
When sufficient CaO was added to make the molar ratios of Ca to Al in the solution equal to 1.8e2.0, 65e72% Al could be recovered from the solution within 4 h. Subsequently, an evaporation step was carried out at 90 C. In this step, NaOH concentration was increased up to 450 g/L allowing the NaCl precipitation, and consequently 92e93% Cl was removed from the electrolyte. The concentrated electrolyte was recycled to the leaching. 5. Conclusion A new approach was suggested to treat flue dusts formed during secondary copper smelting, with the aim of maximizing metals recovery and minimizing generated residues. The method developed for processing the flue dust comprises several steps: Leaching of dusts using a 5 M NaOH solution at 100 g/L S/L ratio and at 80 C for about 1.5 h. These conditions allow a high recovery of Zn and Pb, while minimizing the co-extraction of Cu. Electrolysis of leach liquor at current densities of 100e250 A/m2 which allows the recovery of 83e87.5% Pb from the solution. Pulse current electrolysis of the result solution at Ton ¼ 15 ms and Toff ¼ 10 ms to obtain homogeneous zinc powders, and a recovery of 88e92% of Zn. Precipitation of Cl and Al when their concentrations exceeded 25 g/L and 7 g/L, respectively, which avoids the accumulation of Al and Cl when reusing the solution in the alkaline leaching stage. Integration of alkaline media leaching and two-stage electrolysis can make the whole process economically attractive due to the value-added by-products and environmental friendly due to the nearly-closed cycle. In particular, pulse current makes the zinc powder more uniform and finer, thus improving its marketability. Furthermore, this product could be an attractive material for organic synthesis, brass manufacture and the battery industry. This integration also allows a high level Cl in the electrolyte, simplifying the procedure. The proposed process might contribute to a more sustainable and cost-effective management for other industrial wastes such as hot galvanizing slag, spent catalyst and automobile shredder scrap, but these areas still deserve further attention. Moreover, future research is also needed that explores ways to reduce NaOH consumption and leaching temperature.
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Please cite this article in press as: Qiang, L., et al., Sequential stepwise recovery of selected metals from flue dusts of secondary copper smelting, Journal of Cleaner Production (2014), http://dx.doi.org/10.1016/j.jclepro.2014.03.085