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Comprehensive recovery of metals from cyanidation tailing
Minerals Engineering 70 (2015) 141–147
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Comprehensive recovery of metals from cyanidation tailing C.C. Lv a,b, J. Ding a,b, P. Qian a, Q.C. Li a, S.F. Ye a,⇑, Y.F. Chen a a b
Institute of Process Engineering, Chinese Academy of Sciences, Beijing 100190, China University of Chinese Academy of Sciences, Beijing 100049, China
a r t i c l e
i n f o
Article history: Received 8 June 2014 Accepted 11 September 2014
Keywords: Comprehensive recovery Cyanidation tailing Sodium hypochlorite Cyanide Metals
a b s t r a c t Cyanidation tailing is the residue produced in gold plants which use cyanidation to extract gold. It can be used as a secondary resource to recover residual metals that are of great economic value. The cyanidation tailing investigated in this paper was obtained from Shandong Province, China. It contained valuable metals such as chalcopyrite, galena, sphalerite and pyrite. In this study, alkaline sodium hypochlorite was used as a regulator in the pretreatment stage. It was proved that the sodium hypochlorite played two roles in the flotation pulp: oxidant and pH regulator. On one hand, sodium hypochlorite oxidized cyanide to cyanate, eliminating the negative effect of residual cyanide towards the environment. On the other hand, with the pH of flotation pulp exceeding 10, sphalerite and pyrite were depressed enormously, which was beneficial to the recovery of chalcopyrite and galena. With the Cu–Pb bulk flotation flowsheet, the cyanidation tailing was processed to obtain qualified Cu concentrate with grade of 13.17% and recovery of 70.00% compared with the original Cu grade of 0.21%. The Cu–Pb tailing was processed to obtain qualified Zn concentrate with grade of 34.72% and recovery of 69.58% compared with the original Zn grade of 0.33%, constituting the comprehensive recovery routing for the cyanidation tailing. Ó 2014 Elsevier Ltd. All rights reserved.
1. Introduction Tailings, generated from mineral processing, used to be regarded as deserted byproducts. However, with the exploitation and gradual depletion of the mineral resources all over the world, the reuse of tailings for recycling their residual valuable minerals become a necessity. Cyanidation tailing is the residue produced from gold plants which use cyanidation (direct cyanide leaching process) to extract gold. In China, It is estimated that more than 2.45 million tons of cyanidation tailings are discharged into tailing ponds every year (Zhu et al., 2010). Most of these cyanidation tailings contain some valuable minerals, such as copper minerals, lead minerals, zinc minerals and sulfide minerals. For example, the cyanidation tailing from Penglai Gold Smelting Plant (in Shandong Province, China) contains 38.07% Fe, 41.22% S. This tailing was used to prepare nano-iron red oxide pigment by an ammonia process with urea as precipitant (Li et al., 2008). The cyanidation tailing from Yindongpo Gold Plant (in Henan Province, China) contains 0.14% Cu, 6.40% Pb, 2.83% Zn, 33.06% Fe and those valuable minerals were recovered by flotation using pretreatment method and YO reagent (He et al., 2003). The cyanidation tailing from Tianshui Gold Plant (in Gansu Province, China) contains 1.94% Cu, 5.96% ⇑ Corresponding author. Tel.: +86 13911828468. E-mail address: [email protected] (S.F. Ye). http://dx.doi.org/10.1016/j.mineng.2014.09.007 0892-6875/Ó 2014 Elsevier Ltd. All rights reserved.
Pb, 0.27% Zn, 24.62% Fe. With the combination of a new activator and copper sulfate in the flotation, qualified Cu concentrate can be acquired (Gao and Li, 2005). However, the residual cyanide in cyanidation tailing depresses the recovery of copper minerals, lead minerals, zinc minerals and sulfide minerals (Popov et al., 1988; Grano et al., 1990; Cao and Liu, 2006). Besides, the redundant cyanide is highly toxic to the environment and a threat to human health if not well treated (Shifrin et al., 1996; Mudder and Botz, 2004; Botz et al., 2005; Donato et al., 2007). So the elimination of cyanide (including free cyanide and cyanide complexes) is imperative before retrieving valuable minerals. There are many physical and chemical methods that have been tested for the removal of cyanide from gold mill effluents, including natural degradation, ozonation, bacterial oxidation, SO2 process, acidification, ion exchange, hydrogen peroxide, alkaline chlorination, which are commonly used in treating waste water in gold plants (Botz, 2001; Ritcey, 2005; Kuyucak and Akcil, 2013). Khodadad et al. used sodium and calcium hypochlorite to oxidize cyanide to cyanate and the residual cyanide was under the detection limit under the optimal conditions. They indicated that the removal of cyanide by sodium hypochlorite cost less compared with by calcium hypochlorite, considering operational factors such as transportation and conditioning tanks (Khodadad et al., 2008). Ingles and Scott also pointed out that the oxidation
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of sodium hypochlorite was accepted as a technique for cyanide waste water treatment (Ingles and Scott, 1993). However, all these cyanide destruction methods were used to treat cyanide-containing liquid wastes rather than to treat the cyanide-containing solid wastes. In this study, sodium hypochlorite was chosen as an adjuster to treat the solid waste of cyanidation tailing considering the oxidability and the oxidation condition of sodium hypochlorite. The feed for this study was the cyanidation tailing obtained from Shandong Province in China. It contained some valuable minerals, such as chalcopyrite, galena, sphalerite, and pyrite, so it can be used as a secondary resource to recover these minerals. However, the cyanidation tailing had the following features, bringing about much difficulty to the recovery of valuable minerals: (1) The grades of chalcopyrite, galena, and sphalerite in the tailing were very low. (2) Residual cyanide and reagents in the tailing depressed the recovery of chalcopyrite, galena, and sphalerite to some extent. (3) The particle size was very fine, detrimental to flotation process. 2. Experimental 2.1. Ores characterization The cyanidation tailing in this study was stored in stout polyethylene containers in the dark to minimize the loss of cyanide. The solid content measured was 20% averagely. After filtration, approximately 100 g of sample was dried at about 90 °C in drying oven for 6 h. Then the lumps were broken and thoroughly mixed. Next, the sample was analyzed by chemical analysis methods using atomic absorption spectrophotometer (AAS, WFX-130A, Beijing, China). The result is shown in Table 1. The mineralogical determination of the feed sample was detected by X-ray diffraction (XRD, Smartlab-201307, Rigaku Corporation, Japan) and scanning electron microscope (SEM, JSM-7001F, JEOL, Japan, operating at 15 kV), as shown in Fig. 1. From the chemical composition analysis and the XRD result, it can be seen that the pyrite and quartz were the major minerals in the cyanidation tailing. The valuable metals were mainly copper mineral (grade 0.21%), lead mineral (grade 0.33%), zinc mineral (grade 0.35%) and sulfide mineral (grade 24.98%), which mainly existed in the form of chalcopyrite, galena, sphalerite, separately. The gangues were mainly quartz and silicate. Thus, the goal of the study was to recover Cu, Pb, Zn, and S (representing chalcopyrite, galena, sphalerite and pyrite, the same below) while depressing quartz and silicate. From Fig. 1(b), it can be seen that the minerals were mostly in the form of particle, sheet, and irregular granular. For the particle size analysis, the feed was sieved into different size fractions by wet sieving. Table 2 shows the particle size distribution and the chemical grades of Cu, Pb, Zn, and S in the different size fractions. It was shown in Table 2 that nearly 99% of the cyanidation tailing was less than 74 lm, indicating that the tailing was already very fine and there was no need to further grind it. Besides, Cu, Pb and Zn distributed relatively equally among different size fractions with similar grades. What is more, there were 78.66% of Cu, 80.95% of Pb, and 81.92% of Zn in the 30 lm size fraction. The results showed that these metals cannot be separated by
elutriation to different size fractions, and the particle size were also a big problem for the treatment of cyanidation tailing. It is believed that fine particles typically show slow recovery rates, and are prone to entrainment, which have a considerable impact on grades and recoveries of valuable minerals (Trahar, 1981; Feng and Aldrich, 1999; Graeme, 2012). As a result, the recovery of valuable metals from cyanidation tailing would be of great difficulty. 2.2. Flotation reagents and flowsheet The flotation experiments were conducted with iso-butyl ethionine (Z200), sodium diethyldithiocarbamate (SN9), and sodium nbutyl xanthate (SNBX) as collectors, which were acquired from Jiangxi Copper Corporation. These collectors were all in industrial grade and were diluted in water to a concentration of 1%. Sodium hypochlorite solution (purchased from Sinopharm Chemical Reagent Beijing Co., Ltd.) was added to the pulp as a regulator. Copper sulfate pentahydrate (purchased from Sinopharm Chemical Reagent Beijing Co., Ltd.), in analytical pure, was used to activate sphalerite in the flotation of sphalerite. CP mixture (composed by sulfurous acid, sodium silicate, and carboxy methylated cellulose), was prepared in laboratory. Terpenic oil (acquired from Jiangxi Copper Corporation), with monohydric alcohol content more than 40%, was used in the whole flowsheet as frother. The flotation flowsheet is shown in Fig. 2, following the principle of Cu–Pb bulk flotation. In the Cu–Pb flotation stage, sodium hypochlorite was used as an adjuster, followed by SN9 and Z200 as collectors. At the optimal conditions, Cu–Pb concentrate was collected. Then ultrasonic device and CP mixture was used to separate the Cu–Pb mixture to acquire Cu concentrate and Cu tailing. The Cu–Pb tailing served as the Zn raw material. With the activation of copper sulfate pentahydrate, Zn concentrate can be collected. Zn tailing served as S concentrate. The keypoint would be how to get the Cu concentrate and Zn concentrate from cyanidation tailing in this study. The following tests were designed in the Cu–Pb bulk flotation stage by choosing the primary affecting factors. The effect of sodium hypochlorite dosage, reaction time and the synergism of collectors on the flotation of Cu–Pb bulk flotation were studied through one roughing flotation test. A 300 g of cyanidation tailing sample was transferred into a 1L XFD series single flotation cell (Jilin Exploring Machinery Plant, Jilin, China) and pulped to 30 wt.% solids using city water. It was agitated at 1220 rpm for 1 min before any reagents were added. Then add sodium hypochlorite solution into the pulp for a certain time, followed by SN9 and Z200. As there were residual reagents from former leaching process, there was no need to add frother in the roughing flotation. Thereafter, the pulp was aerated and flotation was carried out. Froth height was maintained at the same level by adding water periodically throughout the test. The bulk concentrate and tailing collected were filtered, dried, and weighed, and the samples were analyzed by chemical analytical method. The results analysis was done with the recoveries of Cu, Pb, Zn and S. 3. Results and discussion 3.1. Influence of sodium hypochlorite on the flotation As shown in Fig. 3, about 83% of Zn was collected without the pretreatment of sodium hypochlorite, while the recoveries of
Table 1 Result of chemical composition analysis of the cyanidation tailing. Elements
Au
Ag
Cu
Pb
Zn
S
As
Fe
Mg
Ca
Al
Si
Content/%
1.10 g/t
18.81 g/t
0.21
0.33
0.35
24.98
0.03
25.08
2.19
2.19
6.50
16.77
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Fig. 1. The XRD (a) and SEM (b) of the cyanidation tailing sample.
Table 2 Elutriation characterization of the feed sample. Size fraction/lm
+74 +50–74 +30–50 +10–30 10 Feed
Occupancy/%
Grade/%
1.05 9.58 8.65 42.77 37.95 100.00
Share/%
Cu
Pb
Zn
S
Cu
Pb
Zn
S
0.20 0.29 0.18 0.22 0.19 0.21
0.02 0.37 0.32 0.31 0.36 0.33
0.00 0.32 0.37 0.38 0.32 0.35
5.56 33.91 32.85 28.40 17.62 24.98
1.00 13.08 7.25 44.71 33.95 100.00
0.06 10.66 8.32 39.87 41.08 100.00
0.00 8.84 9.23 46.89 35.03 100.00
0.23 13.00 11.37 48.62 26.77 100.00
cyanidation tailing
Cu-Pb bulk flotation
Cu-Pb concentrate
Cu-Pb separation
Cu concentrate
Cu-Pb tailing
Zn flotation
Cu tailing
Pb raw material
Zn concentrate
Zn tailing
S concentrate
Fig. 3. Influence of the dosage of sodium hypochlorite on the recovery of metals.
Fig. 2. The sketchy flotation flowsheet of the cyanidation tailing.
aimed Cu and Pb were only 58% and 74%, respectively. When the flotation pulp was treated with 8 kg/t sodium hypochlorite, the recovery of Zn was obviously decreased to 26% and it continuously decreased with the increase of sodium hypochlorite. S had the same tendency with Zn, but with much less recovery. The recoveries of Cu and Pb increased with the increase of sodium hypochlorite from 0 kg/t to 8 kg/t, and then decrease, except that Cu recovery fell back to 64%. Taking all the four metals into consideration, the 8 kg/t of sodium hypoclorite was the optimal dosage, with the recoveries of Cu, Pb, Zn, and S being 65%, 81%, 26% and 9%, respectively.
This phenomenon can be explained by two assumptions. On one hand, the residual cyanide were oxidized by sodium hypochlorite, as shown in Fig. 4. The total cyanide in the pulp sharply decreased from 0.37 g/t to 0.14 g/t with the increase of sodium hypoclorite from 0 kg/t to 16 kg/t. The oxidation and reduction reactions between cyanide and hypochlorite are as follows (Wild et al., 1994; Botz et al., 2005):
CN þ OCl þ H2 O ! CNCl þ 2OH
2OH þ CNCl
CN þ OCl
! CNO þ Cl þ H2 O
! CNO þ Cl
ð1Þ ð2Þ ð3Þ
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hypochlorite led further to higher alkalinity in the pulp, as shown in Fig. 4. The OH- in alkaline pulp combined with the surface Zn2+ and Fe2+ to form hydrophilic membrane, which excluded the adsorption of collector, thus deeply depressed Zn and S to float to the foam layer (Zhu and Zhu, 1996). 3.2. Influence of the reaction time
Fig. 4. Influence of sodium hypochlorite on the cyanide content and pH of pulp.
The effect of reaction time between sodium hypochlorite and the flotation was shown in Fig. 5. It was apparent that the recoveries of Cu and Pb increased with the reaction time (up to 15 min), reaching the maximal recoveries of 86% and 79% respectively, while Zn and S were kept at a low recoveries of 19% and 9% respectively. With longer reaction time (more than 15 min), the recoveries of Cu and Pb both decreased, while the variation of the recoveries of Zn and S were not significant. It indicated that the sodium hypochlorite gradually released ClO- and OH-, and the 15 min was the most appropriate reaction time. 3.3. Influence of synergetic collectors
Fig. 5. Influence of reaction time on the recovery of metals.
The reaction showed the metals’ surface was encircled by much less cyanide. Accordingly the collectors had more chance to combine with the surface positive metal ions in Cu–Pb bulk flotation. On the other hand, the reaction between cyanide and sodium
In the Cu–Pb bulk concentrate, Z200 and SN9 were used as combinated collectors. The metals’ recoveries as a function of the dosage of collectors were shown in Fig. 6. Throughout the range of collector compositions, the recoveries of Cu and Pb followed similar trends. In Fig. 6(a), the dosage of SN9 was fixed at 80 g/t. The recoveries of Cu and Pb firstly increased and then decreased with the dosage of Z200 increasing from 48 g/t to 192 g/t and reached the maximum of 86% and 79% respectively at Z200 of 96 g/t. This was probably because the dosage of Z200 influenced the flotation speed of Cu, Pb and Zn. In Fig. 6(b), the dosage of Z200 was fixed at 96 g/t. Both the recoveries of Cu and Pb increased with the addition of SN9. But when the dosage of SN9 was more than 80 g/t, the recoveries of Cu and Pb only increased slightly, while Zn was floated enormously, which was unfavorable to the Cu–Pb bulk flotation. Thus the optimal dosages of Z200 and SN9 were 96 g/t and 80 g/t, respectively. Z200 is a dialkyl thionocarbamate collector with O-isopropyl-Nethyl thionocarbamate as its major component. It is proved to be selective for chalcopyrite flotation, particularly against gangue iron sulfides (Fairthorne et al., 1996). SN9 is also a typical collector for galena flotation with eminent selectivity (Hu et al., 2004). Mangalam and Khangaonkar used Zeta-potential and adsorption experiments to evaluate the chalcopyrite-SN9 system and proved
Fig. 6. Influence of Z200 (a) and SN9 (b) dosages on the recovery of metals.
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cyanidation tailing 15min 3min
sodium hypoclorite
15kg/t
Z200 SN9
96g/t 80g/t
Cu-Pb rough 3min
Z200 SN 9
48g/t 40g/t
scavenging
concentrate
concentrate
Cu-Pb concentrate
8min
copper sulfate
500g/t
3min
SNBX
300g/t
2min
terpenic oil
20g/t
Zn rough 5min 3min
concentrate
copper sulfate
100g/t
SNBX
100g/t
scavenging concentrate
tailings
Zn concentrate
Fig. 7. Schematic presentation of the closed-circuit flotation.
Table 3 Result of closed-circuit for the closed-circuit flotation. Product
Cu–Pb concentrate Zn concentrate Tailing Sum
Yield/%
4.90 0.82 94.28 100.0
Grade/%
Recovery/%
Cu
Pb
Zn
S
Cu
Pb
Zn
S
4.14 1.06 0.00 0.21
4.86 0.67 0.09 0.33
0.31 34.72 0.06 0.35
14.68 25.36 25.51 24.98
95.87 4.13 0.00 100.0
73.09 1.70 25.20 100.0
4.29 80.68 15.03 100.0
2.88 0.84 96.28 100.0
that SN9 could also be a good collector for chalcopyrite (Mangalam and Khangaonkar, 1985). So these two collectors were compounded so as to enhance the Cu–Pb bulk flotation, as shown in Fig. 6. Many works have been done to prove that the synergism of flotation collectors has a positive effect on the flotation behavior, among which Bradshaw et al. pointed out that using mixtures of collectors led to greater extent of adsorption on the mineral surface, either enhancing the overall hydrophobicity of the mineral surface or resulting in an adsorbed surface layer of collector molecules more suitable for flotation (Bradshaw et al., 1998).
3.5. The closed-circuit process of Cu–Pb bulk flotation Based on the optimal conditions: oxidant sodium hypochlorite of 8 kg/t, reaction time of 15 min, the collector Z200 of 96 g/t and
cyanidation tailing ultrasonic processing
3.4. The flotation of Zn In the collection of Zn, copper sulfate was used to activate sphalerite, then SNBX was added to the flotation system as a collector, obtaining qualified Zn concentrate. It was believed that the copper sulfate reacted with the surface of Zn, forming a layer of CuS, which is more prone to react with SNBX. It caused Zn easier to float than without the activation of copper sulfate (Zhu and Zhu, 1996; Lascelles et al., 2001).
Cu concentrate
5min
CP mixture
3min
Z200
1min
terpenic
3000g/t 20g/t
oil
10g/t
Cu tailing
Fig. 8. Schematic presentation of the separation of Cu–Pb concentrate.
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Table 4 Result of the separation of Cu–Pb concentrate. Product
Yield/%
Cu concentrate Cu tailing Sum
24.12 75.88 100.00
Grade/%
Recovery/%
Cu
Pb
Zn
S
Cu
Pb
Zn
S
13.01 1.32 4.14
3.78 5.20 4.86
0.53 0.24 0.31
13.75 14.97 14.68
75.80 24.20 100.00
18.77 81.23 100.00
41.24 58.76 100.00
22.60 77.40 100.00
SN9 of 80 g/t, the Cu–Pb bulk flotation was done with one roughing-two concentration-one scavenging flowsheet to gain the Cu– Pb bulk concentrate and Cu–Pb tailing. Cu–Pb tailing was treated with copper sulfate of 500 g/t, SNBX of 300 g/t, terpenic oil of 20 g/t, with one roughing-two concentration-one scavenging flowsheet to gain the qualified Zn concentrate, as shown in Fig. 7. Table 3 displayed the result of the closed-circuit process. Table 3 indicated that Cu and Pb were mainly recovered in the Cu–Pb bulk concentrate with recoveries of 95.87% and 73.09%, respectively. Zn was mainly recovered in the Zn concentrate, with Zn grade of 34.72% and recovery of 80.68%. S was recovered in the tailing, with S grade of 25.51% and recovery of 96.28%. 4. Cu–Pb separation The Cu–Pb concentrate was gained after several steps of flotation, and its surface was affected by residual flotation reagents, such as Z200, SN9 and terpenic oil, reducing the difference between Cu and Pb. The key step would be eliminating the residual flotation reagents. By using the ultrasonic device, the ultrasonic wave initiated cavitation in the flotation pulp, cleaning particle surfaces though collapse of cavity bubbles, exposing the fresh surface of particle (Zhou et al., 2009; Zeng and Liang, 2010). After the surface of Cu–Pb concentrate was cleaned, add the CP mixture into the Cu–Pb bulk concentrate pulp to separate Cu and Pb with Z200 as collector, obtaining the Cu concentrate and Cu tailing (Gu, 1992; Liu, 1981). The schematic process was shown in Fig. 8. The separation result was displayed in Table 4. In the Cu–Pb separation, the Cu concentrate got a yield of 24.12%, with Cu grade of 13.01% and Cu recovery of 75.80% as a qualified Cu concentrate, while Pb was recovered mainly in the Cu tailing, with Pb grade of 5.2% and Pb recovery of 81.23%, which can be used to produce Pb concentrate in future research. 5. Conclusions (1) In this flotation work, the effect of sodium hypochlorite on the separation of valuable metals from the cyanidation tailing was analyzed. Sodium hypochlorite played a significant role of oxidizing agent and pH regulator. It oxidized the residual cyanide in cyanidation tailing pulp from 0.37 g/t to 0.14 g/t under test conditions, leading to easier flotation of chalcopyrite and galena than sphalerite and pyrite. (2) The obtaining Cu–Pb concentrate became much harder to separate than the original as a result of the residual reagents. By combining the ultrasonic method with CP mixture depressant, chalcopyrite was separated from galena, obtaining the Cu concentrate with Cu grade of 13.01% and overall Cu recovery of 72.67%. Cu tailing can be used as Pb raw material. Compared with the original Cu grade of 0.21%, it is evidently a big progress. (3) Using the copper sulfate to activate sphalerite was effective in the experiment condition, obtaining Zn concentrate with Zn grade of 34.72% and Zn recovery of 80.68% from original Zn grade of 0.33% in the cyanidation tailing. The Zn tailing could serve as S concentrate for sulfuric acid.
(4) This study indicates that sodium hypochlorite can be an applicable flotation reagent in treating tailings that come from direct cyanide leaching. The low grade metals can be processed to high-quality concentrate that are of great economic value. Besides, the decrease of cyanide in the flotation pulp leads to much less harm to the environment.
Acknowledgements This work was strongly supported by Natural Science Foundation of China (No. 51202249), the 863 Project (2011AA06A104), and Projects in the National Science & Technology Pillar Program during the 12th Five-year Plan Period (2011BAC06B01, 2012BAB08B04). The authors thank all the members in the Material Chemistry and Engineering Group, Institute of Process Engineering, Chinese Academy of Sciences.
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Minerals Engineering journal homepage: www.elsevier.com/locate/mineng
Comprehensive recovery of metals from cyanidation tailing C.C. Lv a,b, J. Ding a,b, P. Qian a, Q.C. Li a, S.F. Ye a,⇑, Y.F. Chen a a b
Institute of Process Engineering, Chinese Academy of Sciences, Beijing 100190, China University of Chinese Academy of Sciences, Beijing 100049, China
a r t i c l e
i n f o
Article history: Received 8 June 2014 Accepted 11 September 2014
Keywords: Comprehensive recovery Cyanidation tailing Sodium hypochlorite Cyanide Metals
a b s t r a c t Cyanidation tailing is the residue produced in gold plants which use cyanidation to extract gold. It can be used as a secondary resource to recover residual metals that are of great economic value. The cyanidation tailing investigated in this paper was obtained from Shandong Province, China. It contained valuable metals such as chalcopyrite, galena, sphalerite and pyrite. In this study, alkaline sodium hypochlorite was used as a regulator in the pretreatment stage. It was proved that the sodium hypochlorite played two roles in the flotation pulp: oxidant and pH regulator. On one hand, sodium hypochlorite oxidized cyanide to cyanate, eliminating the negative effect of residual cyanide towards the environment. On the other hand, with the pH of flotation pulp exceeding 10, sphalerite and pyrite were depressed enormously, which was beneficial to the recovery of chalcopyrite and galena. With the Cu–Pb bulk flotation flowsheet, the cyanidation tailing was processed to obtain qualified Cu concentrate with grade of 13.17% and recovery of 70.00% compared with the original Cu grade of 0.21%. The Cu–Pb tailing was processed to obtain qualified Zn concentrate with grade of 34.72% and recovery of 69.58% compared with the original Zn grade of 0.33%, constituting the comprehensive recovery routing for the cyanidation tailing. Ó 2014 Elsevier Ltd. All rights reserved.
1. Introduction Tailings, generated from mineral processing, used to be regarded as deserted byproducts. However, with the exploitation and gradual depletion of the mineral resources all over the world, the reuse of tailings for recycling their residual valuable minerals become a necessity. Cyanidation tailing is the residue produced from gold plants which use cyanidation (direct cyanide leaching process) to extract gold. In China, It is estimated that more than 2.45 million tons of cyanidation tailings are discharged into tailing ponds every year (Zhu et al., 2010). Most of these cyanidation tailings contain some valuable minerals, such as copper minerals, lead minerals, zinc minerals and sulfide minerals. For example, the cyanidation tailing from Penglai Gold Smelting Plant (in Shandong Province, China) contains 38.07% Fe, 41.22% S. This tailing was used to prepare nano-iron red oxide pigment by an ammonia process with urea as precipitant (Li et al., 2008). The cyanidation tailing from Yindongpo Gold Plant (in Henan Province, China) contains 0.14% Cu, 6.40% Pb, 2.83% Zn, 33.06% Fe and those valuable minerals were recovered by flotation using pretreatment method and YO reagent (He et al., 2003). The cyanidation tailing from Tianshui Gold Plant (in Gansu Province, China) contains 1.94% Cu, 5.96% ⇑ Corresponding author. Tel.: +86 13911828468. E-mail address: [email protected] (S.F. Ye). http://dx.doi.org/10.1016/j.mineng.2014.09.007 0892-6875/Ó 2014 Elsevier Ltd. All rights reserved.
Pb, 0.27% Zn, 24.62% Fe. With the combination of a new activator and copper sulfate in the flotation, qualified Cu concentrate can be acquired (Gao and Li, 2005). However, the residual cyanide in cyanidation tailing depresses the recovery of copper minerals, lead minerals, zinc minerals and sulfide minerals (Popov et al., 1988; Grano et al., 1990; Cao and Liu, 2006). Besides, the redundant cyanide is highly toxic to the environment and a threat to human health if not well treated (Shifrin et al., 1996; Mudder and Botz, 2004; Botz et al., 2005; Donato et al., 2007). So the elimination of cyanide (including free cyanide and cyanide complexes) is imperative before retrieving valuable minerals. There are many physical and chemical methods that have been tested for the removal of cyanide from gold mill effluents, including natural degradation, ozonation, bacterial oxidation, SO2 process, acidification, ion exchange, hydrogen peroxide, alkaline chlorination, which are commonly used in treating waste water in gold plants (Botz, 2001; Ritcey, 2005; Kuyucak and Akcil, 2013). Khodadad et al. used sodium and calcium hypochlorite to oxidize cyanide to cyanate and the residual cyanide was under the detection limit under the optimal conditions. They indicated that the removal of cyanide by sodium hypochlorite cost less compared with by calcium hypochlorite, considering operational factors such as transportation and conditioning tanks (Khodadad et al., 2008). Ingles and Scott also pointed out that the oxidation
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of sodium hypochlorite was accepted as a technique for cyanide waste water treatment (Ingles and Scott, 1993). However, all these cyanide destruction methods were used to treat cyanide-containing liquid wastes rather than to treat the cyanide-containing solid wastes. In this study, sodium hypochlorite was chosen as an adjuster to treat the solid waste of cyanidation tailing considering the oxidability and the oxidation condition of sodium hypochlorite. The feed for this study was the cyanidation tailing obtained from Shandong Province in China. It contained some valuable minerals, such as chalcopyrite, galena, sphalerite, and pyrite, so it can be used as a secondary resource to recover these minerals. However, the cyanidation tailing had the following features, bringing about much difficulty to the recovery of valuable minerals: (1) The grades of chalcopyrite, galena, and sphalerite in the tailing were very low. (2) Residual cyanide and reagents in the tailing depressed the recovery of chalcopyrite, galena, and sphalerite to some extent. (3) The particle size was very fine, detrimental to flotation process. 2. Experimental 2.1. Ores characterization The cyanidation tailing in this study was stored in stout polyethylene containers in the dark to minimize the loss of cyanide. The solid content measured was 20% averagely. After filtration, approximately 100 g of sample was dried at about 90 °C in drying oven for 6 h. Then the lumps were broken and thoroughly mixed. Next, the sample was analyzed by chemical analysis methods using atomic absorption spectrophotometer (AAS, WFX-130A, Beijing, China). The result is shown in Table 1. The mineralogical determination of the feed sample was detected by X-ray diffraction (XRD, Smartlab-201307, Rigaku Corporation, Japan) and scanning electron microscope (SEM, JSM-7001F, JEOL, Japan, operating at 15 kV), as shown in Fig. 1. From the chemical composition analysis and the XRD result, it can be seen that the pyrite and quartz were the major minerals in the cyanidation tailing. The valuable metals were mainly copper mineral (grade 0.21%), lead mineral (grade 0.33%), zinc mineral (grade 0.35%) and sulfide mineral (grade 24.98%), which mainly existed in the form of chalcopyrite, galena, sphalerite, separately. The gangues were mainly quartz and silicate. Thus, the goal of the study was to recover Cu, Pb, Zn, and S (representing chalcopyrite, galena, sphalerite and pyrite, the same below) while depressing quartz and silicate. From Fig. 1(b), it can be seen that the minerals were mostly in the form of particle, sheet, and irregular granular. For the particle size analysis, the feed was sieved into different size fractions by wet sieving. Table 2 shows the particle size distribution and the chemical grades of Cu, Pb, Zn, and S in the different size fractions. It was shown in Table 2 that nearly 99% of the cyanidation tailing was less than 74 lm, indicating that the tailing was already very fine and there was no need to further grind it. Besides, Cu, Pb and Zn distributed relatively equally among different size fractions with similar grades. What is more, there were 78.66% of Cu, 80.95% of Pb, and 81.92% of Zn in the 30 lm size fraction. The results showed that these metals cannot be separated by
elutriation to different size fractions, and the particle size were also a big problem for the treatment of cyanidation tailing. It is believed that fine particles typically show slow recovery rates, and are prone to entrainment, which have a considerable impact on grades and recoveries of valuable minerals (Trahar, 1981; Feng and Aldrich, 1999; Graeme, 2012). As a result, the recovery of valuable metals from cyanidation tailing would be of great difficulty. 2.2. Flotation reagents and flowsheet The flotation experiments were conducted with iso-butyl ethionine (Z200), sodium diethyldithiocarbamate (SN9), and sodium nbutyl xanthate (SNBX) as collectors, which were acquired from Jiangxi Copper Corporation. These collectors were all in industrial grade and were diluted in water to a concentration of 1%. Sodium hypochlorite solution (purchased from Sinopharm Chemical Reagent Beijing Co., Ltd.) was added to the pulp as a regulator. Copper sulfate pentahydrate (purchased from Sinopharm Chemical Reagent Beijing Co., Ltd.), in analytical pure, was used to activate sphalerite in the flotation of sphalerite. CP mixture (composed by sulfurous acid, sodium silicate, and carboxy methylated cellulose), was prepared in laboratory. Terpenic oil (acquired from Jiangxi Copper Corporation), with monohydric alcohol content more than 40%, was used in the whole flowsheet as frother. The flotation flowsheet is shown in Fig. 2, following the principle of Cu–Pb bulk flotation. In the Cu–Pb flotation stage, sodium hypochlorite was used as an adjuster, followed by SN9 and Z200 as collectors. At the optimal conditions, Cu–Pb concentrate was collected. Then ultrasonic device and CP mixture was used to separate the Cu–Pb mixture to acquire Cu concentrate and Cu tailing. The Cu–Pb tailing served as the Zn raw material. With the activation of copper sulfate pentahydrate, Zn concentrate can be collected. Zn tailing served as S concentrate. The keypoint would be how to get the Cu concentrate and Zn concentrate from cyanidation tailing in this study. The following tests were designed in the Cu–Pb bulk flotation stage by choosing the primary affecting factors. The effect of sodium hypochlorite dosage, reaction time and the synergism of collectors on the flotation of Cu–Pb bulk flotation were studied through one roughing flotation test. A 300 g of cyanidation tailing sample was transferred into a 1L XFD series single flotation cell (Jilin Exploring Machinery Plant, Jilin, China) and pulped to 30 wt.% solids using city water. It was agitated at 1220 rpm for 1 min before any reagents were added. Then add sodium hypochlorite solution into the pulp for a certain time, followed by SN9 and Z200. As there were residual reagents from former leaching process, there was no need to add frother in the roughing flotation. Thereafter, the pulp was aerated and flotation was carried out. Froth height was maintained at the same level by adding water periodically throughout the test. The bulk concentrate and tailing collected were filtered, dried, and weighed, and the samples were analyzed by chemical analytical method. The results analysis was done with the recoveries of Cu, Pb, Zn and S. 3. Results and discussion 3.1. Influence of sodium hypochlorite on the flotation As shown in Fig. 3, about 83% of Zn was collected without the pretreatment of sodium hypochlorite, while the recoveries of
Table 1 Result of chemical composition analysis of the cyanidation tailing. Elements
Au
Ag
Cu
Pb
Zn
S
As
Fe
Mg
Ca
Al
Si
Content/%
1.10 g/t
18.81 g/t
0.21
0.33
0.35
24.98
0.03
25.08
2.19
2.19
6.50
16.77
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Fig. 1. The XRD (a) and SEM (b) of the cyanidation tailing sample.
Table 2 Elutriation characterization of the feed sample. Size fraction/lm
+74 +50–74 +30–50 +10–30 10 Feed
Occupancy/%
Grade/%
1.05 9.58 8.65 42.77 37.95 100.00
Share/%
Cu
Pb
Zn
S
Cu
Pb
Zn
S
0.20 0.29 0.18 0.22 0.19 0.21
0.02 0.37 0.32 0.31 0.36 0.33
0.00 0.32 0.37 0.38 0.32 0.35
5.56 33.91 32.85 28.40 17.62 24.98
1.00 13.08 7.25 44.71 33.95 100.00
0.06 10.66 8.32 39.87 41.08 100.00
0.00 8.84 9.23 46.89 35.03 100.00
0.23 13.00 11.37 48.62 26.77 100.00
cyanidation tailing
Cu-Pb bulk flotation
Cu-Pb concentrate
Cu-Pb separation
Cu concentrate
Cu-Pb tailing
Zn flotation
Cu tailing
Pb raw material
Zn concentrate
Zn tailing
S concentrate
Fig. 3. Influence of the dosage of sodium hypochlorite on the recovery of metals.
Fig. 2. The sketchy flotation flowsheet of the cyanidation tailing.
aimed Cu and Pb were only 58% and 74%, respectively. When the flotation pulp was treated with 8 kg/t sodium hypochlorite, the recovery of Zn was obviously decreased to 26% and it continuously decreased with the increase of sodium hypochlorite. S had the same tendency with Zn, but with much less recovery. The recoveries of Cu and Pb increased with the increase of sodium hypochlorite from 0 kg/t to 8 kg/t, and then decrease, except that Cu recovery fell back to 64%. Taking all the four metals into consideration, the 8 kg/t of sodium hypoclorite was the optimal dosage, with the recoveries of Cu, Pb, Zn, and S being 65%, 81%, 26% and 9%, respectively.
This phenomenon can be explained by two assumptions. On one hand, the residual cyanide were oxidized by sodium hypochlorite, as shown in Fig. 4. The total cyanide in the pulp sharply decreased from 0.37 g/t to 0.14 g/t with the increase of sodium hypoclorite from 0 kg/t to 16 kg/t. The oxidation and reduction reactions between cyanide and hypochlorite are as follows (Wild et al., 1994; Botz et al., 2005):
CN þ OCl þ H2 O ! CNCl þ 2OH
2OH þ CNCl
CN þ OCl
! CNO þ Cl þ H2 O
! CNO þ Cl
ð1Þ ð2Þ ð3Þ
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hypochlorite led further to higher alkalinity in the pulp, as shown in Fig. 4. The OH- in alkaline pulp combined with the surface Zn2+ and Fe2+ to form hydrophilic membrane, which excluded the adsorption of collector, thus deeply depressed Zn and S to float to the foam layer (Zhu and Zhu, 1996). 3.2. Influence of the reaction time
Fig. 4. Influence of sodium hypochlorite on the cyanide content and pH of pulp.
The effect of reaction time between sodium hypochlorite and the flotation was shown in Fig. 5. It was apparent that the recoveries of Cu and Pb increased with the reaction time (up to 15 min), reaching the maximal recoveries of 86% and 79% respectively, while Zn and S were kept at a low recoveries of 19% and 9% respectively. With longer reaction time (more than 15 min), the recoveries of Cu and Pb both decreased, while the variation of the recoveries of Zn and S were not significant. It indicated that the sodium hypochlorite gradually released ClO- and OH-, and the 15 min was the most appropriate reaction time. 3.3. Influence of synergetic collectors
Fig. 5. Influence of reaction time on the recovery of metals.
The reaction showed the metals’ surface was encircled by much less cyanide. Accordingly the collectors had more chance to combine with the surface positive metal ions in Cu–Pb bulk flotation. On the other hand, the reaction between cyanide and sodium
In the Cu–Pb bulk concentrate, Z200 and SN9 were used as combinated collectors. The metals’ recoveries as a function of the dosage of collectors were shown in Fig. 6. Throughout the range of collector compositions, the recoveries of Cu and Pb followed similar trends. In Fig. 6(a), the dosage of SN9 was fixed at 80 g/t. The recoveries of Cu and Pb firstly increased and then decreased with the dosage of Z200 increasing from 48 g/t to 192 g/t and reached the maximum of 86% and 79% respectively at Z200 of 96 g/t. This was probably because the dosage of Z200 influenced the flotation speed of Cu, Pb and Zn. In Fig. 6(b), the dosage of Z200 was fixed at 96 g/t. Both the recoveries of Cu and Pb increased with the addition of SN9. But when the dosage of SN9 was more than 80 g/t, the recoveries of Cu and Pb only increased slightly, while Zn was floated enormously, which was unfavorable to the Cu–Pb bulk flotation. Thus the optimal dosages of Z200 and SN9 were 96 g/t and 80 g/t, respectively. Z200 is a dialkyl thionocarbamate collector with O-isopropyl-Nethyl thionocarbamate as its major component. It is proved to be selective for chalcopyrite flotation, particularly against gangue iron sulfides (Fairthorne et al., 1996). SN9 is also a typical collector for galena flotation with eminent selectivity (Hu et al., 2004). Mangalam and Khangaonkar used Zeta-potential and adsorption experiments to evaluate the chalcopyrite-SN9 system and proved
Fig. 6. Influence of Z200 (a) and SN9 (b) dosages on the recovery of metals.
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cyanidation tailing 15min 3min
sodium hypoclorite
15kg/t
Z200 SN9
96g/t 80g/t
Cu-Pb rough 3min
Z200 SN 9
48g/t 40g/t
scavenging
concentrate
concentrate
Cu-Pb concentrate
8min
copper sulfate
500g/t
3min
SNBX
300g/t
2min
terpenic oil
20g/t
Zn rough 5min 3min
concentrate
copper sulfate
100g/t
SNBX
100g/t
scavenging concentrate
tailings
Zn concentrate
Fig. 7. Schematic presentation of the closed-circuit flotation.
Table 3 Result of closed-circuit for the closed-circuit flotation. Product
Cu–Pb concentrate Zn concentrate Tailing Sum
Yield/%
4.90 0.82 94.28 100.0
Grade/%
Recovery/%
Cu
Pb
Zn
S
Cu
Pb
Zn
S
4.14 1.06 0.00 0.21
4.86 0.67 0.09 0.33
0.31 34.72 0.06 0.35
14.68 25.36 25.51 24.98
95.87 4.13 0.00 100.0
73.09 1.70 25.20 100.0
4.29 80.68 15.03 100.0
2.88 0.84 96.28 100.0
that SN9 could also be a good collector for chalcopyrite (Mangalam and Khangaonkar, 1985). So these two collectors were compounded so as to enhance the Cu–Pb bulk flotation, as shown in Fig. 6. Many works have been done to prove that the synergism of flotation collectors has a positive effect on the flotation behavior, among which Bradshaw et al. pointed out that using mixtures of collectors led to greater extent of adsorption on the mineral surface, either enhancing the overall hydrophobicity of the mineral surface or resulting in an adsorbed surface layer of collector molecules more suitable for flotation (Bradshaw et al., 1998).
3.5. The closed-circuit process of Cu–Pb bulk flotation Based on the optimal conditions: oxidant sodium hypochlorite of 8 kg/t, reaction time of 15 min, the collector Z200 of 96 g/t and
cyanidation tailing ultrasonic processing
3.4. The flotation of Zn In the collection of Zn, copper sulfate was used to activate sphalerite, then SNBX was added to the flotation system as a collector, obtaining qualified Zn concentrate. It was believed that the copper sulfate reacted with the surface of Zn, forming a layer of CuS, which is more prone to react with SNBX. It caused Zn easier to float than without the activation of copper sulfate (Zhu and Zhu, 1996; Lascelles et al., 2001).
Cu concentrate
5min
CP mixture
3min
Z200
1min
terpenic
3000g/t 20g/t
oil
10g/t
Cu tailing
Fig. 8. Schematic presentation of the separation of Cu–Pb concentrate.
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Table 4 Result of the separation of Cu–Pb concentrate. Product
Yield/%
Cu concentrate Cu tailing Sum
24.12 75.88 100.00
Grade/%
Recovery/%
Cu
Pb
Zn
S
Cu
Pb
Zn
S
13.01 1.32 4.14
3.78 5.20 4.86
0.53 0.24 0.31
13.75 14.97 14.68
75.80 24.20 100.00
18.77 81.23 100.00
41.24 58.76 100.00
22.60 77.40 100.00
SN9 of 80 g/t, the Cu–Pb bulk flotation was done with one roughing-two concentration-one scavenging flowsheet to gain the Cu– Pb bulk concentrate and Cu–Pb tailing. Cu–Pb tailing was treated with copper sulfate of 500 g/t, SNBX of 300 g/t, terpenic oil of 20 g/t, with one roughing-two concentration-one scavenging flowsheet to gain the qualified Zn concentrate, as shown in Fig. 7. Table 3 displayed the result of the closed-circuit process. Table 3 indicated that Cu and Pb were mainly recovered in the Cu–Pb bulk concentrate with recoveries of 95.87% and 73.09%, respectively. Zn was mainly recovered in the Zn concentrate, with Zn grade of 34.72% and recovery of 80.68%. S was recovered in the tailing, with S grade of 25.51% and recovery of 96.28%. 4. Cu–Pb separation The Cu–Pb concentrate was gained after several steps of flotation, and its surface was affected by residual flotation reagents, such as Z200, SN9 and terpenic oil, reducing the difference between Cu and Pb. The key step would be eliminating the residual flotation reagents. By using the ultrasonic device, the ultrasonic wave initiated cavitation in the flotation pulp, cleaning particle surfaces though collapse of cavity bubbles, exposing the fresh surface of particle (Zhou et al., 2009; Zeng and Liang, 2010). After the surface of Cu–Pb concentrate was cleaned, add the CP mixture into the Cu–Pb bulk concentrate pulp to separate Cu and Pb with Z200 as collector, obtaining the Cu concentrate and Cu tailing (Gu, 1992; Liu, 1981). The schematic process was shown in Fig. 8. The separation result was displayed in Table 4. In the Cu–Pb separation, the Cu concentrate got a yield of 24.12%, with Cu grade of 13.01% and Cu recovery of 75.80% as a qualified Cu concentrate, while Pb was recovered mainly in the Cu tailing, with Pb grade of 5.2% and Pb recovery of 81.23%, which can be used to produce Pb concentrate in future research. 5. Conclusions (1) In this flotation work, the effect of sodium hypochlorite on the separation of valuable metals from the cyanidation tailing was analyzed. Sodium hypochlorite played a significant role of oxidizing agent and pH regulator. It oxidized the residual cyanide in cyanidation tailing pulp from 0.37 g/t to 0.14 g/t under test conditions, leading to easier flotation of chalcopyrite and galena than sphalerite and pyrite. (2) The obtaining Cu–Pb concentrate became much harder to separate than the original as a result of the residual reagents. By combining the ultrasonic method with CP mixture depressant, chalcopyrite was separated from galena, obtaining the Cu concentrate with Cu grade of 13.01% and overall Cu recovery of 72.67%. Cu tailing can be used as Pb raw material. Compared with the original Cu grade of 0.21%, it is evidently a big progress. (3) Using the copper sulfate to activate sphalerite was effective in the experiment condition, obtaining Zn concentrate with Zn grade of 34.72% and Zn recovery of 80.68% from original Zn grade of 0.33% in the cyanidation tailing. The Zn tailing could serve as S concentrate for sulfuric acid.
(4) This study indicates that sodium hypochlorite can be an applicable flotation reagent in treating tailings that come from direct cyanide leaching. The low grade metals can be processed to high-quality concentrate that are of great economic value. Besides, the decrease of cyanide in the flotation pulp leads to much less harm to the environment.
Acknowledgements This work was strongly supported by Natural Science Foundation of China (No. 51202249), the 863 Project (2011AA06A104), and Projects in the National Science & Technology Pillar Program during the 12th Five-year Plan Period (2011BAC06B01, 2012BAB08B04). The authors thank all the members in the Material Chemistry and Engineering Group, Institute of Process Engineering, Chinese Academy of Sciences.
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