metals Article
Recovery of Gold and Iron from Cyanide Tailings with a Combined Direct Reduction Roasting and Leaching Process Pingfeng Fu 1,2, * 1 2
*
ID
, Zhenyu Li 1 , Jie Feng 1 and Zhenzhong Bian 1
School of Civil and Resources Engineering, University of Science and Technology Beijing, Beijing 100083, China;
[email protected] (Z.L.);
[email protected] (J.F.);
[email protected] (Z.B.) Key Laboratory of High-Efficient Mining and Safety of Metal Mines, Ministry of Education, Beijing 100083, China Correspondence:
[email protected]; Tel.: +86-10-6233-2902
Received: 25 June 2018; Accepted: 19 July 2018; Published: 21 July 2018
Abstract: Cyanide tailings are the hazardous waste discharged after gold cyanidation leaching. The recovery of gold and iron from cyanide tailings was investigated with a combined direct reduction roasting and leaching process. The effects of reduction temperature, coal dosage and CaO dosage on gold enrichment into Au-Fe alloy (Fex Au1−x ) were studied in direct reduction roasting. Gold containing iron powders, i.e., Au-Fe alloy, had the gold grade of 8.23 g/t with a recovery of 97.46%. After separating gold and iron in iron powders with sulfuric acid leaching, ferrous sulfate in the leachate was crystallized to prepare FeSO4 ·7H2 O with a yield of 222.42% to cyanide tailings. Gold enriched in acid-leaching residue with gold grade of 216.58 g/t was extracted into pregnant solution. The total gold recovery of the whole process reached as high as 94.23%. The tailings generated in the magnetic separation of roasted products, with a yield of 51.33% to cyanide tailings, had no toxic cyanide any more. The gold enrichment behaviors indicated that higher reduction temperature and larger dosage of coal and CaO could promote the allocation of more gold in iron phase rather than in slag phase. The mechanism for enriching gold from cyanide tailings into iron phase was proposed. This work provided a novel route to simultaneously recover gold and iron from cyanide tailings. Keywords: cyanide tailings; hazardous waste; direct reduction roasting; Au-Fe alloy; acid leaching; gold recovery
1. Introduction Cyanidation leaching, which discharges a great amount of cyanide tailings, is the predominant process to produce gold around the world [1,2]. As toxic cyanides and heavy metals such as mercury and arsenic remain in cyanide tailings, they are extremely harmful to human beings and animals, and are classified as hazardous wastes in China [3–7]. Nevertheless, the content of valuable metals such as gold and iron may be high in cyanide tailings. The gold grade and total iron (TFe) content of some tailings can reach 3-9 g/t and above 30%, respectively [8–10]. Therefore, attention should be paid to recovering valuable metals from cyanide tailings [11–15]. Since sulfide minerals are still present in cyanide residues discharged after direct cyanidation leaching, flotation is an efficient method to recover noble metals and non-ferrous metals such as Cu, Zn and Pb [16–18]. However, noble metals in cyanide tailings discharged after the roasting-cyanidation leaching are enclosed in iron oxide minerals (e.g., hematite). To recover iron in the tailings, the magnetization roasting-magnetic separation is investigated to produce iron concentrates by reducing
Metals 2018, 8, 561; doi:10.3390/met8070561
www.mdpi.com/journal/metals
Metals 2018, 8, 561
2 of 13
hematite to magnetite. However, some reports showed that the TFe content of iron concentrates was below 60% due to high content of impurities such as SiO2 and Al2 O3 [8]. The acid-leaching ammonia process was also tested to dissolve iron oxides in cyanide tailings with nitric or sulfuric acid and to prepare iron oxide pigments [19]. To recover noble metals in cyanide tailings, magnetization roasting is applied to promote gold extraction, and the gold leaching rate of 46.14% was reported [9]. However, magnetite in roasted products is found to have a negative effect on gold leaching, which leads to low Au leaching rate [20]. Since the gold can react with Cl2 or HCl to generate volatile gold chlorides at high temperature, the chlorination roasting has been studied to recover gold from cyanide tailings with the gold recovery of above 90% [21,22]. However, this process has deficiencies involving Cl2 and HCl gases which can corrode the equipment and is extremely harmful to human health. Up to now, the simultaneous recovery of gold and iron from cyanide tailings has rarely been investigated. Although acid leaching can achieve high gold recovery and produce FeSO4 or iron oxide pigments [19], this process is complex and costly due to large acid consumption and severe leaching conditions. Magnetization roasting can promote the recovery of iron and gold from cyanide tailings, but the gold recovery is quite low [9,20]. Therefore, it is necessary to develop an efficient and feasible process to recover gold and iron from cyanide tailings after the roasting-cyanidation leaching. Atomic Au and Fe can form Au-Fe alloys at high temperature [23]. The unique electrical [24], magnetic [25], thermodynamic [26] and catalytic [27] properties of Au-Fe alloys have attracted great interest. It suggests that gold remains in cyanide tailings can be captured and enriched in metallic iron if a high-temperature reduction process is conducted. The direct reduction-magnetic separation can effectively recover iron from converter slag [28], vanadium tailings [29], copper slag [30,31], and lead smelting slag [32] to produce iron powders. In this work, direct reduction roasting is applied to produce gold containing iron powders, namely Fe-rich Au-Fe alloy (Fex Au1−x ), from cyanide tailings. The gold and iron in iron powders are separated by sulfuric acid leaching. Then the leachate is used to produce FeSO4 ·7H2 O crystals with a low temperature crystallization method. Although the extraction of gold with alternative lixiviants (e.g., thiosulfate and bromide) do not produce cyanide residue [33,34], the cyanidation extraction is conducted to recover gold from acid-leaching residue because generated cyanide residue can be recycled as raw materials into direct reduction roasting, thus no secondary cyanide residue is produced in this process. The objectives of this study are: (1) to evaluate the technical feasibility of simultaneous recovery of gold and iron from cyanide tailings via the combined direct reduction roasting and leaching process; (2) to investigate the effects of reduction temperature, coal dosage and CaO dosage on the enrichment of gold into Au-Fe alloy in direct reduction roasting; and (3) to propose the mechanism for gold enrichment from cyanide tailings into iron phase. The results in this work can provide a novel route to effectively recover gold and iron from cyanide tailings, pyrite cinders and gold containing iron ores. 2. Materials and Methods 2.1. Materials and Reagents The cyanide tailings discharged after the roasting-cyanidation leaching were received from a gold smelting plant in Henan Province, China. The main chemical composition was given in Table 1. The TFe content and gold grade reached 44.96% and 3.83 g/t, respectively, indicating high content of valuable metals. The mineral composition revealed by X-ray powder diffraction (XRD) (Figure 1) indicated that dominant minerals were hematite and quartz. The coal with a size of less than 0.5 mm, used as the reducing agent, had 9.81% of ash, 26.34% of volatiles, 0.73% of moisture and 63.12% of fixed carbon based on its industrial analysis. The CaO in analytical grade was purchased from Sinopharm Chemical Reagent Co., Ltd. The CaO content was higher than 98.0% as given by the supplier.
Metals 2018, 8, 561
3 of 13
Table 1. Main chemical composition of received cyanide tailings. Metals 2018, 8, x FOR PEER REVIEW
3 of 14
Metals 2018,Components 8, x FOR PEER REVIEW TFe
Au * SiO2 Al2 O3 CaO MgO Na2 O K2 O Components TFe Au* SiO2 Al2O3 CaO MgO Na2O K2O Content (%) (%) 44.9644.96 2.47 0.58 MgO 0.67 Na Content 3.83 24.35 2.47 0.58 0.67 1.2621.26 Components TFe 3.83 Au* 24.35 SiO2 Al 2O3 CaO O 1.15 K2O1.15 * * The24.35 unit of Au was The unitg/t. of0.58 Au was0.67 g/t. Content (%) 44.96 3.83 2.47 1.26 1.15 *
1
The unit of Au was g/t.
1- hematite 21- quartz hematite
1
(a.u.) Intensity (a.u.) Intensity
3 of 14
1
2- quartz
1 2
2
1 1
1 2
1
1
1
2 1
2
2 50
1
10
2 20
30
2 40
10
20
30
40
11 1 11 1 2 1 1 2 60 70
2θ50(°)
60
1 11
1 801 1 90
70
80
90
(°) cyanide tailings. Figure 1. XRD pattern of2θ received
Figure 1. XRD pattern of received cyanide tailings. Figure 1. XRD pattern of received cyanide tailings.
2.2. Direct Reduction Roasting of Cyanide Tailings
2.2. Direct Reduction Roasting of Cyanide Tailings
The recovery was shown in Figure 2. Cyanide tailings were dried at 105 °C overnight. 2.2. Direct Reductionprocess Roasting of Cyanide Tailings ◦ C overnight. The process was shown Figure Cyanide tailings were dried at 105 30 g recovery tailings were thoroughly mixedinwith coal2.and CaO with a designated dosage. Herein, the The recovery process was shown in Figure 2. Cyanide tailings were dried at 105 °C overnight. 30 g tailings were with by coala and CaO with designated Herein, the dosage dosage of coalthoroughly or CaO wasmixed expressed percentage thatareferred to itsdosage. mass ratio to cyanide 30 g tailings were thoroughly mixed with coal and CaO with a designated dosage. Herein, the tailings. The was dosages of coal and were chosen in the range from wt.% ratio to 20 wt.% and from 5 of coal or CaO expressed by CaO a percentage that referred to its8.0mass to cyanide tailings. dosage of coal or CaO was expressed by a percentage that referred to its mass ratio to cyanide wt.% to 20 wt.%, respectively. The mixture was added to a graphite crucible with a lip. The direct The dosages of coal and CaO were chosen in the range from 8.0 wt.% to 20 wt.% and from 5 wt.% to tailings. The dosages of coal and CaO were chosen in the range from 8.0 wt.% to 20 wt.% and from 5 reduction roasting The was mixture conducted in aadded muffletofurnace (CD-1400X, Zhengzhou Chida Tungsten & 20 wt.%, was aadded graphite with a lip. The direct wt.% respectively. to 20 wt.%, respectively. The mixture was to a crucible graphite crucible with a lip. The reduction direct Molybdenum Products Co. Ltd., Zhengzhou, China). When the reduction temperature reached its roasting was conducted in conducted a muffle furnace (CD-1400X, Zhengzhou Chida Tungsten &Tungsten Molybdenum reduction in acrucible muffle furnace (CD-1400X, Zhengzhou Chida designatedroasting value ofwas 1050‒1250 °C, the was transferred into the furnace and maintained for& Products Co. Ltd.,Products Zhengzhou, China). When the reduction temperature reached its designated value Molybdenum Co. Ltd., Zhengzhou, China). When the reduction temperature reached 90 min at◦ a reduction atmosphere. After that, the crucible was taken out of the furnace and cooledits of 1050–1250 C, the crucible was transferred into the furnace and maintained for 90 min at a reduction designated value of 1050‒1250 °C, the crucible was transferred into the furnace and maintained for down to room temperature in air. atmosphere. that, the crucible was taken out ofcrucible the furnace cooled to room temperature 90 min atAfter a reduction atmosphere. After that, the was and taken out ofdown the furnace and cooled down to room temperature in air. in air. cyanide tailings 3.83 100 Gold grade /g/t Yield /%
TFe content /% Gold grade /g/t Yield /%
TFe content /%
44.96
3.83 100 cyanide tailings Coal + CaO 44.96
direct reduction Coal + CaO
direct reduction magnetic separation 8.23 45.34 0.19 51.33 magnetic separation iron powders tailings 92.43 5.94
8.23 45.34 acid leachingtailings 0.19 51.33 iron powders 92.43 5.94 216.58 1.72 leaching residue acid leaching 28.71 leachate 216.58 1.72 cyanidation leaching residue 28.71 extraction
leachate crystallization
7.10 1.72 cyanidation / pregnant cyanide extraction
7.10 1.72
residue
crystallization / 222.42 solution FeSO4·7H2O 19.49
/ 222.42 cyanide of pregnant Figure 2. Flowsheet for/ the recovery gold and FeSO iron from cyanide tailings at the reduction 4·7H2O 19.49 solution residue temperature of 1200 °C. (The Yield is defined as the weight percentage of solid products to received cyanide2.tailings as shown in Equation Figure Flowsheet for the recovery(5)). of gold and iron from cyanide tailings at the reduction
Figure 2. Flowsheet for the recovery of gold and iron from cyanide tailings at the reduction temperature temperature of 1200 °C. (The Yield is defined as the weight percentage of solid products to received of 1200 ◦ C. (The Yield is defined as the weight percentage of solid products to received cyanide tailings cyanide tailings as shown in Equation (5)). as shown in Equation (5)).
Metals 2018, 8, 561
4 of 13
The roasted product was firstly ground to lower than 0.074 mm fraction of 70.2% in a rod mill (RK/BM, Wuhan Rock Crush & Grand Equipment Manufacture Co., Ltd., Wuhan, China), then separated by a magnetic tube separator (XCGS-73, Xinsheng Mining Equipment Co., Ltd., Jiangxi, China) at a field intensity of 167.06 kA/m. The concentrate was further ground to lower than 0.045 mm fraction of 80.5% and separated at 107.41 kA/m to obtain final iron concentrate, i.e., reduced iron powders containing gold. The tailings generated in two stage magnetic separation as shown in Figure 2 were collected as magnetic separation tailings. The TFe or gold recovery of reduced iron powders was calculated as follows: W × β1 ε= 1 × 100% (1) W0 × β 0 where, ε was the TFe or gold recovery (%), W 1 was the weight of reduced iron powders (g), β1 was the TFe content (%) or gold grade (g/t) of iron powders, W 0 was the weight of cyanide tailings (g), and β0 was the TFe content (%) or gold grade (g/t) of cyanide tailings. 2.3. Sulfuric Acid Leaching and Preparation of Ferrous Sulfate Crystals 10 g reduced iron powders were mixed with 140 mL of 2.35 mol/L H2 SO4 solution in a conical flask. The acid leaching was performed at different temperatures for 1 h in a thermostatic bath. After the leaching, the pulp was filtered to obtain leaching residue with enriched gold. The hot leachate was firstly condensed by heating at 100 ◦ C. After cooling down to room temperature, the condensed leachate was kept in a refrigerator at 0 ◦ C for 12 h to crystallize FeSO4 ·7H2 O. Herein, the weight loss ratio (η) and iron leaching rate (γ), calculated as Equations (2) and (3), were used to evaluate acid-leaching process, respectively. W1 − W3 × 100% W1
(2)
W1 × β 1 − W3 × β 3 × 100% W1 × β 1
(3)
η= γ=
where, η and γ were the weight loss ratio (%) and iron leaching rate (%), respectively, W 3 was the weight of acid-leaching residue (g), β3 was the TFe content (%) of acid-leaching residue. 2.4. Cyanidation Extraction of Gold from Acid-Leaching Residue The gold enriched in acid-leaching residue was finally recovered by cyanidation extraction. The extraction was conducted in a conical flask without air bubbling and stirred with a magnetic stirrer at 300 rpm for 36 h. The liquid to solid mass ratio was set to 10:1, and pulp pH was 11.0. The cyanide dosage was in a range of 5–80 kg/t. After the leaching, the pulp was filtered to obtain pregnant solution. The cyanide residue was dried to assay gold grade. The Au extraction rate (φ) was calculated as follows: W3 × β 3 − W5 × β 5 φ= × 100% (4) W3 × β 3 where, φ was the gold extraction rate (%), β3 and β5 was the gold grade (g/t) of acid-leaching residue and cyanide residue, respectively, W 5 was the weight (g) of cyanide residue. In this process, the yield (δi ) of solid products to cyanide tailings in Figure 1 was defined as Equation (5): W δi = i × 100% (5) W0 where, W i (i.e., W 0 , W 1 , W 2 , W 3 , W 4 and W 5 ) was the weight (g) of cyanide tailings, iron powders, magnetic separation tailings, acid-leaching residue, FeSO4 ·7H2 O crystals and cyanide residue, respectively.
Metals 2018, 8, 561 Metals 2018, 8, x FOR PEER REVIEW
5 of 13 5 of 14
2.5. 2.5. Analysis Analysis and and Characterization Characterization The Augrade gradeof of solid products was determined by inductively plasma-optical The Au solid products was determined by inductively coupled coupled plasma-optical emission emission spectrometry (ICP-OES, iCAP 7000, Thermo Fisher Scientific Inc., Waltham, MA, USA) at spectrometry (ICP-OES, iCAP 7000, Thermo Fisher Scientific Inc., Waltham, MA, USA) at radio radio frequency (RF) power of 1.35 air flow of 14 mL/min,and andargon argonauxiliary auxiliary flow flow rate rate of frequency (RF) power of 1.35 kW, kW, air flow raterate of 14 mL/min, of 0.3 mL/min. 0.3 g samples were digested with 3 mL ionized water, 9 mL HCl and 3 mL HNO 3 at 130 0.3 mL/min. 0.3 g samples were digested with 3 mL ionized water, 9 mL HCl and 3 mL HNO3 °C for 15 closed acid digester Grand Analytical Instrument Co. Ltd., ◦ Cmin at 130 for in 15 amin in amicrowave closed microwave acid (DS-360x, digester (DS-360x, Grand Analytical Instrument Guangzhou, China). The TFe content ofcontent solid products analyzed by a titanium (III) chloride Co. Ltd., Guangzhou, China). The TFe of solid was products was analyzed by a titanium (III) reduction and potassium dichromate titration method (Chinese standard, GB/T 6730.65-2009). The chloride reduction and potassium dichromate titration method (Chinese standard, GB/T 6730.65-2009). contents of SiO Al2O, 3Al , CaO, MgO, Na2O and K2O in cyanide tailings, and of Mn, Pb, As, Si, Al, and The contents of2,SiO 2 2 O3 , CaO, MgO, Na2 O and K2 O in cyanide tailings, and of Mn, Pb, As, Si, Cu ·7H 2O crystals were also obtained by ICP-OES. The mineral compositions of cyanide Al, in andFeSO Cu 4in FeSO 4 ·7H2 O crystals were also obtained by ICP-OES. The mineral compositions of tailings, roasted and acid-leaching residue were investigated by XRD by (DMax-RB, Rigaku cyanide tailings, products roasted products and acid-leaching residue were investigated XRD (DMax-RB, Corporation, Tokyo, Japan) conditions of 40 kV, 100 The Rigaku Corporation, Tokyo, under Japan) the under the conditions of 40 kV,mA 100and mACu andKα Curadiation. Kα radiation. morphology of roasted products was performed with scanning electron microscopy (SEM, EVO18, The morphology of roasted products was performed with scanning electron microscopy (SEM, EVO18, Carl Corporation, Oberkochen, Oberkochen, Germany). Germany). Roasted Carl Zeiss Zeiss Group Group Corporation, Roasted products products were were mounted mounted on on epoxy epoxy resin and further polished before the SEM observation. resin and further polished before the SEM observation. 3. 3. Results Results and and Discussion Discussion
Recovery of of Gold Gold and and Iron Iron with Direct Reduction Roasting 3.1. Recovery 3.1.1. Effect Effect of of Reduction Reduction Temperature Temperature 3.1.1.
100
100
90
95 90
80 85 70 TFe content TFe recovery Gold recovery
60 50
80 75
Recovery of metal (%)
TFe content (%)
As shown shown in in Figure Figure 3, 3, the the TFe TFe content content and and recovery recovery of of reduced reduced iron iron powders powders were were increased increased As from 68.84% to 92.43% and from 78.98% to 93.19%, respectively, as the temperature was raised from from 68.84% to 92.43% and from 78.98% to 93.19%, respectively, as the temperature was raised from ◦ 1050 to suggests that high temperature can promote the reduction of ironof oxides metallic 1050 to 1200 1200 C. °C.It It suggests that high temperature can promote the reduction iron to oxides to iron. Meanwhile, the gold recovery was increased from 74.33% to 95.03%, indicating the enrichment of metallic iron. Meanwhile, the gold recovery was increased from 74.33% to 95.03%, indicating the ◦ remained gold in the tailings irontailings powders. thepowders. temperature wastemperature raised up to was 1250 raised C, theup gold enrichment of remained goldinto in the intoAs iron As the to recovery was further increased to 95.70%. It clearly indicated that higher temperature could promote 1250 °C, the gold recovery was further increased to 95.70%. It clearly indicated that higher the capture ofcould morepromote gold by the in situ generated metallic iron. temperature capture of more gold by in situ generated metallic iron.
70 1050
1100
1150
1200
1250
Reduction temperature (°C) Figure Figure 3. 3. Effect Effect of of reduction reduction temperature temperature on on the the recovery recovery of of iron iron and and gold gold from from cyanide cyanide tailings. tailings. (Reduction conditions: reduction time of 90 min, coal dosage of 20 wt.% and CaO dosage (Reduction conditions: reduction time of 90 min, coal dosage of 20 wt.% and CaO dosage of of 15 15 wt.%). wt.%).
The direct reduction of hematite to metallic iron generally follows the pathway of Fe2O3→Fe3O4 The direct reduction of hematite to metallic iron generally follows the pathway of Fe2 O3 →Fe3 O4 →FeO→Fe [35]. At above 900 °C, Fe3O4 can be further reduced to FeO. As the reduction of FeO to →FeO→Fe [35]. At above 900 ◦ C, Fe O4 can be further reduced to FeO. As the reduction of FeO to metallic iron is a strong endothermic3 reaction, high temperature is necessary to generate metallic metallic iron is a strong endothermic reaction, high temperature is necessary to generate metallic iron. iron. Due to the existence of SiO2 in cyanide tailings, FeO can react with SiO2 to form fayalite, a low Due to the existence of SiO2 in cyanide tailings, FeO can react with SiO2 to form fayalite, a low melting melting point material promoting to form molten slag [36]. The amount of molten slag increase point material promoting to form molten slag [36]. The amount of molten slag increase sharply at sharply at above 1140 °C, which can facilitate the growth of iron grains and enhance the separation
of iron grains from slag [37,38]. Therefore, as shown in Figure 3, the elevated temperature could increase the TFe content and recovery of iron powders. 3.1.2. Effect of Coal Dosage Metals 2018, 8, 561
6 of 13
As shown in Figure 4, when the coal dosage was increased from 8 wt.% to 20 wt.%, the TFe and gold recovery was rapidly increased from 56.36% to 93.03% and from 42.53% to 93.19%, respectively. above 1140 ◦ C, which can facilitate the growth of iron grains and enhance the separation of iron grains However, the TFe content of iron powders was just increased from 89.92% to 92.43%. When the from slag [37,38]. Therefore, as shown in Figure 3, the elevated temperature could increase the TFe dosage was further raised up to 25 wt.%, the TFe content slightly decreased, but almost no change of content and recovery of iron powders. the TFe and gold recovery was observed. Thus, the optimum coal dosage was 20 wt.%. The result revealed thatofhigher coal dosage could promote to recover gold and iron from cyanide tailings. 3.1.2. Effect Coal Dosage The high coal dosage can increase the C/O (fixed carbon/oxygen) ratio in the reduction system, Asenhances shown inthe Figure 4, when the coal dosage was increased from 8 wt.% to 20 wt.%, the TFe iron and which reduction atmosphere and promotes the reduction of more FeO to metallic gold Thus, recovery rapidly increased from iron 56.36% 93.03% and from 42.53% 93.19%, respectively. [39]. thewas higher amount of metallic cantosurely capture more gold.toThe FeO in reduction However, the TFe content of iron powders was just increased from 89.92% to 92.43%. When the [38,40]. dosage system is a substance decreasing the liquidus temperature by forming a fayalite-type slag was further up to 25 wt.%, TFe content slightly no less, change of the TFe When more raised FeO was reduced to the Fe at higher C/O ratio, decreased, the moltenbut slagalmost became which might and gold recovery was observed. Thus, the optimum coal dosage was 20 wt.%. The result revealed reduce the mechanical mixing of gold and slag in slag phase. Therefore, the allocation of gold in that higher dosage could promote to recover gold and iron from cyanide tailings. metallic ironcoal phase should be enhanced. 100
TFe content (%)
90 95
80 70
90 60 TFe content TFe recovery Gold recovery
85
80 8
12
16
20
50
Recovery of metal (%)
100
40 30 28
24
Coal dosage (wt.%) Figure 4. Effect of coal dosage on the recovery of iron and gold from cyanide tailings. (Reduction Figure 4. Effect of coal dosage on the recovery of iron and gold from cyanide tailings. (Reduction conditions: reduction temperature of 1200 °C, reduction time of 90 min, CaO dosage of 15 wt.%). conditions: reduction temperature of 1200 ◦ C, reduction time of 90 min, CaO dosage of 15 wt.%).
3.1.3. Effect of CaO Dosage The high coal dosage can increase the C/O (fixed carbon/oxygen) ratio in the reduction system, Asenhances shown in 5, when the CaO dosage wasthe increased from 5 wt.% wt.%, iron the [39]. TFe which theFigure reduction atmosphere and promotes reduction of more FeOtoto20 metallic recovery was raised slightly, but the Au recovery significantly was raised from 77.65% to 97.46%. Thus, the higher amount of metallic iron can surely capture more gold. The FeO in reduction system is The TFe content showedthe a week relationship withby CaO dosage. Thus, the optimum CaO When dosagemore was a substance decreasing liquidus temperature forming a fayalite-type slag [38,40]. 20 wt.%. resulttorevealed that the CaO dosage a remarkable effect onwhich the gold recovery. FeO wasThe reduced Fe at higher C/O ratio, the had molten slag became less, might reduce the
TFe content (%)
3.1.3. Effect of CaO Dosage
95
95
Recovery of metal (%)
mechanical mixing of gold and slag in slag phase. Therefore, the allocation of gold in metallic iron 100 100 phase should be enhanced.
As shown in Figure 5, when the CaO dosage was increased from 5 wt.% to 20 wt.%, the TFe 90 recovery was raised slightly, but the Au recovery significantly was raised from 77.65% to 97.46%. 90 The TFe content showed a week relationship with CaO dosage. Thus, the optimum CaO dosage was 85 20 wt.%. The result revealed that the CaO dosage had a remarkable effect on the gold recovery. TFe content TFe recovery Gold recovery
85
80 75
80 4
8
12
16
CaO dosage (wt.%)
20
3.1.3. Effect of CaO Dosage As shown in Figure 5, when the CaO dosage was increased from 5 wt.% to 20 wt.%, the TFe recovery was raised slightly, but the Au recovery significantly was raised from 77.65% to 97.46%. The content MetalsTFe 2018, 8, 561 showed a week relationship with CaO dosage. Thus, the optimum CaO dosage7 was of 13 20 wt.%. The result revealed that the CaO dosage had a remarkable effect on the gold recovery.
TFe content (%)
95
95
90 90 85 TFe content TFe recovery Gold recovery
85
80
Recovery of metal (%)
100
100
75
80 4
8
12
16
20
CaO dosage (wt.%) Figure 5. Effect of CaO dosage on the recovery of iron and gold from cyanide tailings. (Reduction conditions: reduction temperature of 1200 ◦ C, reduction time of 90 min, coal dosage of 20 wt.%).
The SiO2 content in cyanide tailings reached 24.35% as shown in Table 1. Reactive FeO can react with SiO2 to form fayalite at high temperature (Equation (6)), which hinders the reduction of FeO to metallic Fe. While the CaO is added, the fayalite may readily react with fixed carbon and/or CaO to form Fe and wollastonite (Equations (7) and (8)) [41,42], which generates more metallic iron. The results revealed that higher alkalinity (w(CaO)/w(SiO2 )) of slag phase could promote more gold allocated in iron phase rather than in slag phase. SiO2 + 2FeO = Fe2 SiO4 , ∆GTθ = −33231 + 12.26 T
(6)
Fe2 SiO4 + 2C = 2Fe + SiO2 + CO, ∆GTθ = 167700 − 160.91 T
(7)
Fe2 SiO4 + 2C + CaO = 2Fe + CaO·SiO2 + CO, ∆GTθ = 235347 − 310.71 T
(8)
Under the optimum reduction roasting conditions: roasting temperature of 1200 ◦ C, roasting time of 90 min, coal dosage of 20 wt.% and CaO dosage of 20 wt.%, iron powders with TFe content of 92.43% and gold grade of 8.23 g/t were achieved. The TFe and gold recovery in the direct reduction roasting reached 93.21% and 97.46%, respectively. As toxic cyanides were decomposed at 1200 ◦ C, the magnetic separation tailings would not contain cyanides. Thus, magnetic separation tailings had lower toxicity and weight (a yield of 51.33% to cyanide tailings) compared to cyanide tailings. 3.2. Proposed Mechanism for Enrichment of Gold from Cyanide Tailings Figure 6 showed the phase diagram of Au-Fe binary system [23]. Pure Au and γ-Fe crystals have same space groups (Fm3m) and close lattice parameters (0.40784 nm for Au and 0.36468 nm for γ-Fe). As shown in Figure 6, Au and γ-Fe atoms can form Au-rich fcc terminal solid solution (Au) and Fe-rich γ terminal solid solution based on fcc phase of Fe (γ-Fe) at high temperature [23,43]. The solubility limit of Au in the (γ-Fe) phase was reported to be 4.1–8.1 at.% [23]. The liquidus (L) of Au-Fe system has a minimum of 1036 ◦ C and 18.5 at.% of Fe. In this work, when roasted in 1050 ◦ C, remained gold in cyanide tailings may react with in situ generated γ-Fe to form Au-rich (Au) phase or liquidus (L) even the reduction temperature was below the melting point of pure gold (1064.43 ◦ C). As the roasting time was extended, more γ-Fe was generated by the reduction of FeO, promoting the formation of (γ-Fe) phase with low composition of Au. When the roasting temperature was increased over the gold melting point, solid gold particles in the reduction system became molten, increasing the mobility of Au atoms. Thus, these Au atoms distributed in slag phase may be readily transferred into iron phase to form Fe-rich (γ-Fe) phase due to high mobility. Therefore, with the increase of reduction temperature, more gold in cyanide tailings could be allocated into iron phase rather than in slag phase.
Metals 2018, 8, 561
8 of 13
Metals 2018, 8, x FOR PEER REVIEW
8 of 14
Metals 2018, 8, x FOR PEER REVIEW Weight percent of Fe 5
10
15
30
20
40 50 60 70 80 100 1538
L
1400
℃
(δ-Fe) 1538 1433 89 97 1394 (δ-Fe) 1433 89 97 1394 L+(γ-Fe) ℃
℃
℃
℃
1200
1173 L+(γ-Fe)
57 1064.43
1064.43 1000
57
℃
74
18.5% 1038
800
92 868
53
℃
53
(Au)
769
769
℃
10
20
30
40
50
912
98 99912 (Au)+(α-Fe) 868
60
℃
℃
℃
℃
(α-Fe)
(Tc)
0
98 ℃
770
(α-Fe) 99 (Au)+(α-Fe) (Tc)
600 600
(γ-Fe)
℃
(Au)
800
(γ-Fe)
92
℃
℃
18.5% 1038
1000
℃
74 1173
70
80
L 1433
℃
℃
(δ-Fe)
97 98
℃
(δ-Fe)
℃
(γ-Fe) 1394 1400 L+(γ-Fe) 1350 (γ-Fe) 1394 94 96 98 100 1350 Atomic percent of Fe (%) 94 96 98 100 950 Atomic percent of Fe (%) 912 (γ-Fe) 950 900 912 (γ-Fe) 96 900 99 868 850 96 (α-Fe) 99 868 850 (Au)+(α-Fe) 800 (α-Fe) 770 769 (Tc) (Au)+(α-Fe) 800 750 770 769 (Tc) 97 98 99 100 750 Atomic percent of Fe (%) 97 98 99 100 ℃
℃
℃
℃
1200
1500 1450
1538
1450 1400 1433 L+(γ-Fe) 97 98
℃
Temperature (°C) Temperature (°C)
℃
L
1400
1538
L
℃
50 60 70 80 100
℃
1600 0 1600
Temperature ( ) Temperature ( )
5
1550 1550 1500
Temperature ( ) Temperature ( )
0
Weight of Fe 30 40 20 10 15 percent
8 of 14
770
℃
℃
90 100
℃
℃
℃
℃
℃
℃
℃
percent of Fe 0 Au10 20 30Atomic 40 50 60 70 80 90 100Fe Atomic percent of Fe Au Fe Atomic percent of Fe (%) Figure 6. Phase diagram of Au-Fe binary system. Figure 6. Phase diagram of Au-Fe binary system. Figure 6. Phase diagram of Au-Fe binary system.
Figure 7 showed XRD patterns of roasted products obtained from 1050 to 1250 °C. Metallic iron ◦ C. Metallic iron Figure 77 showed patterns of roasted products obtained from 1050 Figure showed XRD patterns of roasted products obtained fromassigned 1050 to to 1250 1250 °C. Metallic iron was generated evenXRD at 1050 °C. The intensity of diffraction peaks to metallic iron became ◦ was generated even at C. The intensity diffraction assigned metallic iron was generated evenhigher at 1050 1050amount °C. Theand intensity of diffraction peaks assigned toelevated metallictemperature. iron became becameAs strong, revealing larger of crystal size ofpeaks metallic iron atto strong, revealing higher amount and larger crystal size metallic at°C elevated strong, revealing higher and crystal of of metallic at elevated shown in Figure 8, theamount grain size oflarger metallic ironsize (grayish white)iron atiron 1050 wastemperature. justtemperature. about 20As µ m, ◦ As shown in Figure 8, the grain size of metallic iron (grayish white) at 1050 C was just about 201200 shown in Figure 8, the grainwere size separated. of metallic Nevertheless, iron (grayish white) at 1050 just about µµm, m, °C and most of these grains some grain size°Cofwas metallic iron at20 ◦C and most of grains were separated. Nevertheless, someiron grainparticles size of of metallic metallic iron at 1200 and most of these theseup grains were100 separated. some grain size iron at 1200 °C was increased to about µ m and Nevertheless, dispersed metallic had been aggregated into was increased up to about 100 µm and dispersed metallic iron particles had been aggregated into was increased up to about 100 µ m and dispersed metallic iron particles had been aggregated into larger grains. Herein, it should point out that metallic iron grains containing gold is recovered by larger grains. ititshould point grains containing gold is recovered byofthe larger grains. Herein, Herein, should pointout outthat thatmetallic metallic iron grains containing gold is recovered by the grinding and magnetic separation process. Theiron larger grain size means easy separation iron grinding and magnetic separation process. The larger grain size means easy separation of iron grains thegrains grinding separation The method larger grain means easy separation of iron as fromand themagnetic slags by the physicalprocess. separation such size as magnetic separation. Therefore, from the slags the3,physical separation method such magnetic separation. Therefore,Therefore, as shown as in grains from thebyslags by the physical separation method such as magnetic separation. shown in Figure the higher Au and Fe recovery ofasiron powders was observed. Figure 3, the higher Au and Fe recovery of iron powders was observed. shown in Figure 3, the higher Au and Fe recovery of iron powders was observed. 1 2
1
2
Intensity (a.u.) Intensity (a.u.)
2
2
2 2 2
2
2
2
2
2
22
2 2 2 4 32 3 2 4 2 3 4 44 32 223 4 3 33 2 3 4 3 2 3 2 4 3 34 32 33 3 3 3 3 3 3 3 3 4 3 4 4 3 3 34 3 3 3 4 34 3 3 4 3 33 4 3 34 3 3 3
1-metallic iron 2-pseudowollastonite 1-metallic iron 3-wollastonite 2-pseudowollastonite 4-fayalite 3-wollastonite 1 4-fayalite 1 1250C1 1 1250C
1200C 1 1200C 1 1 1150C 1 3 3 1 1150C 1 1 3 3 1 1100C 3 3 1 3 3 1 1100C 1 1 1050C 3 3 3 1 1 1050C 3 3
23
1
1
20 30 40 50 60 70 80 90 20 30 40 50 2θ60 70 80 90 (º)
2θ (º)
Figure 7. XRD patterns of roasted products obtained at different roasting temperatures. Figure roasting temperatures. temperatures. Figure 7. 7. XRD XRD patterns patterns of of roasted roasted products products obtained obtained at at different different roasting
Metals 2018, 8, 561 Metals 2018, 8, x FOR PEER REVIEW Metals 2018, 8, x FOR PEER REVIEW
9 of 13 9 of 14 9 of 14
Figure 8. SEM images of roasted products obtained at 1050°C (a) and 1200°C (b) with same Figure 8. SEM images Figure 8. SEM images of of roasted roasted products products obtained obtained at at 1050°C 1050◦ C (a) (a) and and 1200°C 1200◦ C (b) (b) with with same same magnification of 200. magnification of 200. magnification of 200.
100 100
100 100
90 90
90 90
80 80
80 80 Weight loss ratio Weight loss ratio Iron leaching rate Iron leaching rate
70 70 60 60
30 30
40 40
50 50
60 60
70 70
80 80
Leaching temperature (°C) Leaching temperature (°C)
90 90
70 70
Iron (%) rate(%) leachingrate Ironleaching
Weight (%) ratio(%) lossratio Weightloss
As shown in Figure 7, as the roasting temperature was increased, the fayalite, a low melting As shown in Figure 7, as the roasting temperature was increased, the fayalite, a low melting pointAs material, lower 7, fraction slag phase and disappeared at 1250 the °C. fayalite, At the same time, more shown had in Figure as the in roasting temperature was increased, a low melting point material, had lower fraction in slag phase and disappeared at 1250 °C. At the same time, more ◦ wollastonite (α-CaSiO 3 ) in slag phase was transformed to pseudowollastonite (β-CaSiO 3) at point material, had lower fraction in slag phase and disappeared at 1250 C. At the same time, wollastonite (α-CaSiO3) in slag phase was transformed to pseudowollastonite (β-CaSiO3) at temperature higher(α-CaSiO than 1150 °C sincephase the transformation temperature of α-CaSiO3 to (β-CaSiO β-CaSiO3 was more wollastonite in slag was transformed to pseudowollastonite ) at 3 ) °C temperature higher than 1150 since the transformation temperature of α-CaSiO3 to β-CaSiO3 3was 1126 °C. Compared to fayalite of 1205 °C), (pseudo)wollastonite the higher temperature higher than 1150 ◦ C (melting since the point transformation temperature of α-CaSiO3 tohad β-CaSiO was 3 1126 ◦°C. Compared to fayalite (melting point of 1205 °C), (pseudo)wollastonite had the higher ◦ C), (pseudo)wollastonite melting of 1540 °C [35]. (melting Thus, thepoint amount of liquidus in reduction system remarkable 1126 C. point Compared to fayalite of 1205 had would the higher melting melting point◦ of 1540 °C [35]. Thus, the amount of liquidus in reduction system would remarkable decrease as more pseudowollastonite generated at highersystem reduction temperature, preventing point of 1540 C [35]. Thus, the amount was of liquidus in reduction would remarkable decrease as decrease as more pseudowollastonite was generated at higher reduction temperature, preventing the mechanical mixing of molten gold in slag phase. Thus, more gold can be allocated into iron more pseudowollastonite was generated at higher reduction temperature, preventing the mechanical the mechanical mixing of molten gold in slag phase. Thus, more gold can be allocated into iron phase to form Au-Fe It was wellmore in agreement higher at elevated mixing of molten gold inalloy. slag phase. Thus, gold can be with allocated intoAu ironrecovery phase to form Au-Fe phase to form Au-Fe alloy. It was well in agreement with higher Au recovery at elevated temperature (Figure 3) and with increased dosage (Figure 5). temperature (Figure 3) and with alloy. It was well in agreement with higherCaO Au recovery at elevated temperature (Figure 3) and with increased CaO dosage (Figure 5). increased CaO dosage (Figure 5). 3.3. Sulfuric Acid Leaching of Gold Containing Iron Powders 3.3. Sulfuric Acid Leaching of Gold Containing Iron Powders 3.3. Sulfuric Acid Leaching of Gold Containing Iron Powders Sulfuric acid leaching was used to separate gold and iron in iron powders. The effect of leaching Sulfuric acid leaching was used to separate gold and iron in iron powders. The effect of leaching Sulfuric acid leaching was used separate gold and iron iron powders. effect of leaching temperature on iron dissolution was to shown in Figure 9. As theintemperature wasThe raised from 30 to 50 temperature on iron dissolution was shown in Figure 9. As the temperature was raised from 30 to 50 temperature on loss ironratio dissolution was shown rate in Figure As the temperature raised from to °C, the weight and iron leaching were 9.increased from 71.41%was to 96.20% and 30 from °C,◦the weight loss ratio and iron leaching rate were increased from 71.41% to 96.20% and from 50 C, the weight loss ratio andWhile iron leaching rate werewas increased to 96.20% from 82.97% to 98.82%, respectively. the temperature further from raised71.41% to 90 °C, the ironand leaching 82.97% to 98.82%, respectively. While the temperature was further raised to 90 ◦°C, the iron leaching 82.97% to 98.82%, respectively. While the temperature was further to 90 C, the iron leaching rate became almost constant. At acid-leaching temperature of 50raised °C, the yield of acid-leaching rate became almost constant. At acid-leaching temperature of◦ 50 °C, the yield of acid-leaching rate became almost tailings constant.was At acid-leaching temperature 50 grade C, the reached yield of acid-leaching residue residue to cyanide reduced to 1.72%, and theofgold as high as 216.58 g/t residue to cyanide tailings was reduced to 1.72%, and the gold grade reached as high as 216.58 g/t to cyanide was reduced togold 1.72%, and the gold reached as high as 216.58 (Figure 1), (Figure 1), tailings revealing remarkable enrichment andgrade significant weight decrease of g/t acid-leaching (Figure 1), revealing remarkable gold enrichment and significant weight decrease of acid-leaching revealing residue. remarkable gold enrichment and significant weight decrease of acid-leaching residue. residue.
60 60
Figure 9. Effect of acid-leaching temperature on iron dissolution of of reduced reduced iron iron powders. Figure 9. 9. Effect Effect of of acid-leaching acid-leaching temperature temperature on on iron iron dissolution dissolution Figure of reduced iron powders. powders.
Figure 10 showed XRD pattern of acid-leaching residue. The diffraction peaks could be Figure 10 showed XRD pattern of acid-leaching residue. The diffraction peaks could be assigned to cohenite (Fe3C) and bustamite (CaMn(SiO3)2), generated in the direct reduction roasting. assigned to cohenite (Fe3C) and bustamite (CaMn(SiO3)2), generated in the direct reduction roasting.
Metals 2018, 8, 561
10 of 13
showed XRD pattern MetalsFigure 2018, 8, x10FOR PEER REVIEW
of acid-leaching residue. The diffraction peaks could be assigned 10 of 14 to cohenite (Fe3 C) and bustamite (CaMn(SiO3 )2 ), generated in the direct reduction roasting. A broad peak with a 2θwith degree ofdegree 10–40 ◦of C was observed, indicatingindicating the appearance of amorphous materials. A broad peak a 2θ 10–40 °C was observed, the appearance of amorphous However, However, no diffraction peak of metallic iron was observed, complete dissolution materials. no diffraction peak of metallic iron wasrevealing observed,nearly revealing nearly complete ◦ C. Therefore, it can reasonably infer that the Au-Fe of metallic iron in dilute H SO solution at 50 dissolution of metallic iron 2in dilute H2SO4 solution at 50 °C. Therefore, it can reasonably infer that 4 alloyAu-Fe structure iron powders be completely broken, andbroken, gold should be fully exposed in the alloyinstructure in ironmust powders must be completely and gold should be fully acid-leaching residue. In this work, In metallic iron in iron powders to readily react with dilute exposed in acid-leaching residue. this work, metallic iron inwas ironfound powders was found to readily H2 SOwith 50 ◦ C according to Equations (9–11). However,(9–11). sulfuric acid leaching of react dilute HFeSO 2SO4 to produce FeSO4 at 50 °C according to Equations However, sulfuric 4 to produce 4 at ◦ C) and with iron oxides (e.g., hematite) be conducted at be much higher temperature (90-110 acid leaching of iron oxidesshould (e.g., hematite) should conducted at much higher temperature (90‒ longer time (2-4 h) [44,45]. 110°C) and with longer time (2‒4 h) [44,45]. Fe(s) + H2 SO4(l) = FeSO4(l) + H2(g) (9) Fe(s) + H2SO4(l) = FeSO4(l) + H2(g) (9) 4FeSO4(l) + 2H2 SO4(l) + O2(g) = 2Fe2 (SO4 )3(l) + 2H2 O(l) (10) 4FeSO4(l) + 2H2SO4(l) + O2(g) = 2Fe2(SO4)3(l) + 2H2O(l) (10) Fe(s) + Fe2 (SO4 )3(l) = 3FeSO4(l) (11) Fe(s) + Fe2(SO4)3(l) = 3FeSO4(l) (11)
Intensity (a.u.)
2
1-Cohenite 2-Bustamite
1
2 2 2
1
111 1 1 2
10
20
30
40
1 1
50
60
70
80
2θ () Figure Figure 10. 10. XRD XRD pattern pattern of of acid-leaching acid-leaching residue. residue.
The ferrous sulfate in leachate was recovered to produce FeSO4·7H2O crystals with a low The ferrous sulfate in leachate was recovered to produce FeSO4 ·7H2 O crystals with a low temperature crystallization process. The yield of FeSO4·7H2O to cyanide tailings was as high as temperature crystallization process. The yield of FeSO4 ·7H2 O to cyanide tailings was as high as 222.42% as shown in Figure 1. The chemical composition in Table 2 showed that the content of 222.42% as shown in Figure 1. The chemical composition in Table 2 showed that the content of FeSO4·7H2O reached 96.73% with low content of harmful elements such as Pb and As. The obtained FeSO ·7H O reached 96.73% with low content of harmful elements such as Pb and As. The obtained FeSO44·7H22O crystals can be used as raw materials in the production of pigment, water treatment FeSO4 ·7H2 O crystals can be used as raw materials in the production of pigment, water treatment chemicals and feed for animals, etc. chemicals and feed for animals, etc. Table 2. Chemical composition of prepared FeSO4·7H2O crystals. Table 2. Chemical composition of prepared FeSO4 ·7H2 O crystals.
Components Components Content (%)
FeSO4·7H2O FeSO 4 ·7H2 O 96.73
Mn Mn 0.011
Pb
As
Si
Al
Pb 0.002
As 0.0001
Si 0.03
Al 0.01
Cu Cu 0.003
Content (%)
96.73
0.011
0.002
0.0001
0.01
0.003
0.03
3.4. Recovery of Gold from Acid-Leaching Residue
3.4. Recovery of Goldgold frominAcid-Leaching The enriched acid-leachingResidue residue was finally extracted by the cyanidation extraction. As given in Table 3,gold the in gold extraction rate was was increased higher cyanide At the cyanide The enriched acid-leaching residue finallyatextracted by the dosage. cyanidation extraction. dosage ofin40Table kg/t,3,the rate was up to at 96.72%, revealed As given theextraction gold extraction rateincreased was increased higherwhich cyanide dosage.nearly At thecomplete cyanide exposure of gold in acid-leaching residue. The gold in pregnant solution was not further recovered dosage of 40 kg/t, the extraction rate was increased up to 96.72%, which revealed nearly complete at the current stage, and further studies adsorption by solution activatedwas carbon or ion exchange exposure of gold in acid-leaching residue.such The as gold in pregnant not further recovered of at gold wouldstage, be taken. As shown in Figure the gold grade of produced residue reached the current and further studies such as1,adsorption by activated carboncyanide or ion exchange of gold 7.10 g/t,bewhich be recycled thegold direct reduction roasting as raw materials. Therefore, would taken. could As shown in Figureinto 1, the grade of produced cyanide residue reached 7.10 g/t, there was no toxic secondary cyanide residue produced in this process. Table 3. The gold extraction rate of acid-leaching residue.
Metals 2018, 8, 561
11 of 13
which could be recycled into the direct reduction roasting as raw materials. Therefore, there was no toxic secondary cyanide residue produced in this process. Table 3. The gold extraction rate of acid-leaching residue. Cyanide Dosage (kg/t)
5
10
20
40
80
Gold extraction rate (%)
56.78
87.46
94.48
96.72
96.93
In this work, the total gold recovery of whole process, i.e., gold recovery in direct reduction roasting (97.46%) × gold extraction rate (96.72%), was as high as 94.23%, showing high gold recovery of this combined process in treating cyanide tailings. As the gold in acid-leaching residue was fully exposed and had high grade of 216.58 g/t, the cyanide consumption per unit weight of gold was as low as 0.185 kg (NaCN)/g(gold). However, for gold ores (e.g., gold grade of 2 g/t), the cyanide consumption usually reaches 0.5-1 kg (NaCN)/g (gold) for the huge amount of ores although the cyanide dosage (1-2 kg/t) is much lower than that in this work (Table 3) [46]. Thus, it was worthwhile noting that the cyanide consumption in this process was much lower than that in gold extraction industries. 4. Conclusions In this work, the combined direct reduction roasting and leaching process could effectively recover gold and iron from cyanide tailings discharged after the roasting-cyanidation leaching. The effects of direct reduction parameters on the gold enrichment were studied. Gold containing iron powders, i.e., Fe-rich Au-Fe alloy, were produced by the direct reduction roasting-magnetic separation. Under the optimum conditions—reduction temperature of 1200 ◦ C, reduction time of 90 min, coal dosage of 20 wt.% and CaO dosage of 20 wt.%—the TFe content and recovery of iron powders reached 92.43% and 93.21%, respectively, and the gold grade and recovery reached 8.23 g/t and 97.46%, respectively. The magnetic separation tailings, with a yield of 51.33% to cyanide tailings, had no toxic cyanides any more. The gold and iron in iron powders were further separated by sulfuric acid leaching. The leachate was used to prepared FeSO4 ·7H2 O crystals with a yield of 222.42% to cyanide tailings. The acid-leaching residue with the gold grade of 216.58 g/t was subjected to the cyanidation extraction with gold extraction rate of 96.72%. The total gold recovery of whole process was as high as 94.23%. The cyanide residue generated could be recycled into direct reduction roasting as gold containing raw materials. The increase of reduction temperature, coal dosage and CaO dosage resulted in higher Au recovery in iron powders. The higher dosage of coal and CaO promoted the reduction of fayalite to metallic iron and increased the melting temperature of slag phase, which enhanced more gold allocated in iron phase rather than in slag phase. As the mineralogical properties of cyanide tailings after the roasting-cyanidation leaching were close to that of pyrite cinders and gold containing iron ores, the novel process investigated could be applied in valuable metal recovery from these tailings and natural ores. However, in this work, the enrichment of silver in cyanide tailings into iron powders had not been studied. In the future work, the enrichment behaviors of silver in the direct reduction should be fully investigated. Author Contributions: P.F. designed the experiments and prepared the manuscript. Z.L. conducted the XRD and SEM tests. F.J. performed the experiments of direct reduction roasting, magnetic separation and acid leaching. Z.B. conducted the cyanidation extraction and analyzed the element contents of samples. Funding: This research was funded by the National Natural Science Foundation of China, grant number 51674017. Acknowledgments: The authors would like to appreciate the China Scholarship Council for funding the State Scholarship, grant number 201706465070, and thank to Experiment Center in School of Metallurgical and Ecological Engineering, University of Science and Technology Beijing for providing access to analytical facilities. Conflicts of Interest: The authors declare no conflict of interest.
Metals 2018, 8, 561
12 of 13
References 1. 2.
3. 4. 5.
6.
7.
8. 9. 10. 11. 12. 13.
14. 15.
16. 17. 18. 19. 20. 21. 22. 23.
Sayiner, B. Influence of lead nitrate on cyanide leaching of gold and silver from Turkish gold ores. Physicochem. Probl. Miner. Process. 2014, 50, 507–513. [CrossRef] Yazici, E.Y.; Ahlatci, F.; Koc, E.; Celep, O.; Deveci, H. Pre-treatment of a copper-rich gold ore for elimination of copper interference. In Proceedings of the European Metallurgical Conference, Weimar, Germany, 23–26 June 2016. Donato, D.B.; Nichols, O.; Possingham, H.; Moore, M.; Ricci, P.F.; Noller, B.N. A critical review of the effects of gold cyanide-bearing tailings solutions on wildlife. Environ. Int. 2007, 33, 974–984. [CrossRef] [PubMed] Zhong, Q.; Yang, Y.B.; Chen, L.J.; Li, Q.; Xu, B.; Jiang, T. Intensification behavior of mercury ions gold cyanide leaching. Metals 2018, 8, 80. [CrossRef] Clement, A.J.H.; Novakova, T.; Hudson-Edwards, K.A.; Fuller, I.C.; Macklin, M.G.; Fox, E.G.; Zapico, I. The environmental and geomorphological impacts of historical gold mining in the Ohinemuri and Waihou river catchments, Coromandel, New Zealand. Geomorphology 2017, 295, 159–175. [CrossRef] Koohestani, B.; Darban, A.K.; Darezereshki, E.; Mokhtari, P.; Yilmaz, E.; Yilmaz, E. The influence of sodium and sulfate ions on total solidification and encapsulation potential of iron-rich acid mine drainage in silica gel. J. Environ. Chem. Eng. 2018, 6, 3520–3527. [CrossRef] Yilmaz, E.; Ahlatci, F.; Yazici, E.Y.; Celep, O.; Deveci, H. Passivation of pyritic tailings and mitigation of its acid mine drainage (AMD) potential. In Proceedings of the International Symposium on Mining and Environment, Mugla, Turkey, 27–29 September 2017. Zhang, Y.L.; Li, H.M.; Yu, X.J. Recovery of iron from cyanide tailings with reduction roasting-water leaching followed by magnetic separation. J. Hazard. Mater. 2012, 213-214, 167–174. [CrossRef] [PubMed] Liu, B.L.; Zhang, Z.H.; Li, L.B.; Wang, Y.J. Recovery of gold and iron from the cyanide tailings by magnetic roasting. Rare. Metal. Mater. Eng. 2013, 42, 1805–1809. Pil´sniak-Rabiega, M.; Trochimczuk, A.W. Selective recovery of gold on functionalized resins. Hydrometallurgy 2014, 146, 111–118. [CrossRef] De Andrade Lima, L.R.P.; Bernardez, L.A.; Barbosa, L.A.D. Characterization and treatment of artisanal gold mine tailings. J. Hazard. Mater. 2008, 150, 747–753. [CrossRef] [PubMed] Syed, S. Recovery of gold from secondary sources—A review. Hydrometallurgy 2012, 115–116, 30–51. [CrossRef] Koc, E.; Ahlatci, F.; Kuzu, M.; Yazici, E.Y.; Celep, O.; Deveci, H. Recovery of silver from cyanide leach solutions of a pyritic gold concentrate by sodium sulphide precipitation. In Proceedings of the 15th International Mineral Processing Symposium and Exhibition, Istanbul, Turkey, 11–13 June 2016. Yazici, E.Y.; Yilmaz, E.; Ahlatci, F.; Celep, O.; Deveci, H. Recovery of silver from cyanide leach solutions by precipitation using Trimercapto-s-triazine (TMT). Hydrometallurgy 2016, 174, 175–183. [CrossRef] Ilankoon, I.M.S.K.; Tang, Y.; Ghorbani, Y.; Northey, S.; Yellishetty, M.; Deng, X.Y.; McBride, D. The current state and future directions of percolation leaching in the Chinese mining industry: Challenges and opportunities. Miner. Eng. 2018, 125, 206–222. [CrossRef] Dehghani, A.; Ostad-rahimi, M.; Mojtahedzadeh, S.H.; Gharibi, K.K. Recovery of gold from the mouteh gold mine tailings dam. J. South African Inst. Min. Metall. 2009, 109, 417–421. Lv, C.C.; Ding, J.; Qian, P.; Li, Q.C.; Ye, S.F.; Chen, Y.F. Comprehensive recovery of metals from cyanidation tailing. Miner. Eng. 2015, 70, 141–147. [CrossRef] Yang, X.L.; Huang, X.; Qiu, T.S. Recovery of zinc from cyanide tailings by flotation. Miner. Eng. 2015, 84, 100–105. [CrossRef] Li, D.X.; Gao, G.L.; Meng, F.L.; Ji, C. Preparation of nano-iron oxide red pigment powders by use of cyanided tailings. J. Hazard. Mater. 2008, 155, 369–377. [CrossRef] Bas, A.D.; Safizadeh, F.; Ghali, E.; Choi, Y. Leaching and electrochemical dissolution of gold in the presence of iron oxide minerals associated with roasted gold ore. Hydrometallurgy 2016, 166, 143–153. [CrossRef] Li, Z.Y.; Wang, W.W.; Yue, K.; Chen, M.X. High-temperature chlorination of gold with transformation of iron phase. Rare Met. 2016, 35, 881–886. [CrossRef] Li, H.Y.; Zhang, L.B.; Koppala, S.; Ma, A.Y.; Peng, J.H.; Li, S.W.; Yin, S.H. Extraction of gold and silver in the selective chlorination roasting process of cyanidation tailing. Sep. Sci. Technol. 2018, 53, 458–466. [CrossRef] Okamoto, H.; Massalski, T.B.; Swartzendruber, L.J.; Beck, P.A. The Au-Fe (gold-iron) system. Bull. Alloy Phase Diagrams 1984, 5, 592–601. [CrossRef]
Metals 2018, 8, 561
24. 25.
26. 27. 28. 29. 30. 31. 32. 33. 34. 35. 36.
37. 38. 39. 40. 41. 42. 43. 44. 45. 46.
13 of 13
Bansal, C.; Sarkar, S.; Mishra, A.K.; Abraham, T.; Lemier, C.; Hahn, H. Electronically tunable conductivity of a nanoporous Au–Fe alloy. Scr. Mater. 2007, 56, 705–708. [CrossRef] Lee, Y.P.; Kudryavtsev, Y.V.; Nemoshkalenko, V.V.; Gontarz, R.; Rhee, J.Y. Magneto-optical and optical properties of Fe-rich Au-Fe alloy films near the fcc-bcc structural transformation region. Phys. Rev. B 2003, 67, 104424. [CrossRef] Konieczny, R.; Idczak, R.; Chojcan, J. A study of thermodynamic properties of dilute Fe-Au alloys by the 57Fe Mössbauer Spectroscopy. Acta Phys. Pol. A 2017, 131, 255–258. [CrossRef] Kumar, P.S.M.; Sivakumar, T.; Fujita, T.; Jayavel, R.; Abe, H. Synthesis of metastable Au-Fe alloy using ordered nanoporous silica as a hard template. Metals 2018, 8, 17. [CrossRef] Xiang, J.Y.; Huang, Q.Y.; Lv, X.W.; Bai, C.G. Multistage utilization process for the gradient-recovery of V, Fe, and Ti from vanadium-bearing converter slag. J. Hazard. Mater. 2017, 336, 1–7. [CrossRef] [PubMed] Yang, H.F.; Jing, L.L.; Zhang, B.G. Recovery of iron from vanadium tailings with coal-based direct reduction followed by magnetic separation. J. Hazard. Mater. 2011, 185, 1405–1411. [CrossRef] [PubMed] Cao, Z.C.; Sun, T.C.; Xue, X.; Liu, Z.H. Iron recovery from discarded copper slag in a RHF direct reduction and subsequent grinding/magnetic separation process. Minerals 2016, 6, 119. [CrossRef] Zhang, B.J.; Niu, L.P.; Zhang, T.G.; Li, Z.Q.; Zhang, D.L.; Zheng, C. Alternative reduction of copper matte in reduction process of copper slag. ISIJ Int. 2017, 57, 775–781. [CrossRef] Li, W.F.; Zhan, J.; Fan, Y.Q.; Wei, C.; Zhang, C.F.; Hwang, J.Y. Research and industrial application of a process for direct reduction of molten high-lead smelting slag. JOM 2017, 69, 784–789. [CrossRef] Xu, B.; Kong, W.H.; Li, Q.; Yang, Y.B.; Jiang, T.; Liu, X.L. A review of thiosulfate leaching of gold: Focus on thiosulfate consumption and gold recovery from pregnant solution. Metals 2017, 7, 222. [CrossRef] Ahtiainen, R.; Lundstrom, M. Preg-robbing of gold in chloride-bromide solution. Physicochem. Probl. Miner. Process. 2016, 52, 244–251. [CrossRef] Liu, G.S.; Strezov, V.; Lucas, J.A.; Wibberley, L.J. Thermal investigations of direct iron ore reduction with coal. Thermochim. Acta 2004, 410, 133–140. [CrossRef] Li, Y.L.; Sun, T.C.; Kou, J.; Guo, Q.; Xu, C.Y. Study on phosphorus removal of high-phosphorus oolitic hematite by coal-based direct reduction and magnetic separation. Min. Proc. Ext. Met. Rev. 2014, 35, 66–73. [CrossRef] Han, H.; Duan, D.; Yuan, P.; Chen, S. Recovery of metallic iron from high phosphorous oolitic hematite by carbothermic reduction and magnetic separation. Ironmak. Steelmak. 2015, 42, 542–547. [CrossRef] Yu, W.; Tang, Q.Y.; Chen, J.A.; Sun, T.C. Thermodynamic analysis of the carbothermic reduction of a high-phosphorus oolitic iron ore by FactSage. Int. J. Min., Met. Mater. 2016, 23, 1126–1132. [CrossRef] Yu, W.; Sun, T.C.; Cui, Q.; Xu, C.Y.; Kou, J. Effect of coal type on the reduction and magnetic separation of a high-phosphorus oolitic hematite ore. ISIJ Int. 2015, 55, 536–543. [CrossRef] Zhou, X.L.; Zhu, D.Q.; Pan, J.; Wu, T.J. Utilization of waste copper slag to produce directly reduced iron for weathering resistant steel. ISIJ Int. 2015, 55, 1347–1352. [CrossRef] Cheng, X.L.; Zhao, K.; Qi, Y.H.; Shi, X.F.; Zheng, C.L. Direct reduction experiment on iron-bearing waste slag. J. Iron Steel Res. Int. 2013, 20, 24–29. [CrossRef] Mousa, E.A. Effect of basicity on wustite sinter reducibility under simulated blast furnace conditions. Ironmak. Steelmak. 2014, 41, 418–429. [CrossRef] Favez, D.; Wagnière, J.D.; Rappaz, M. Au-Fe alloy solidification and solid-state transformations. Acta Mater. 2010, 58, 1016–1025. [CrossRef] Liu, Z.R.; Zeng, K.; Zhao, W.; Li, Y. Effect of temperature on iron leaching from bauxite residue by sulfuric acid. Bull. Environ. Contam. Toxicol. 2009, 82, 55–58. [CrossRef] [PubMed] Zheng, Y.J.; Liu, Z.C. Preparation of monodispersed micaceous iron oxide pigment from pyrite cinders. Powder Technol. 2011, 207, 335–342. [CrossRef] Bas, A.D.; Koc, E.; Yazici, E.Y.; Deveci, H. Treatment of copper-rich gold ore by cyanide leaching, ammonia pretreatment and ammoniacal cyanide leaching. Trans. Nonferrous Met. Soc. China 2015, 25, 597–607. [CrossRef] © 2018 by the authors. Licensee MDPI, Basel, Switzerland. This article is an open access article distributed under the terms and conditions of the Creative Commons Attribution (CC BY) license (http://creativecommons.org/licenses/by/4.0/).
