Recovery of Gold and Iron from Cyanide Tailings with a Combined Direct Reduction Roasting and Leaching Process

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Article Recovery of Gold and Iron from Cyanide Tailings with a Combined Direct Reduction Roasting and Leaching Process

Pingfeng Fu 1,2,* ID , Zhenyu Li 1, Jie Feng 1 and Zhenzhong Bian 1

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.) 2 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 (FexAu1−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 FeSO4·7H2O 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 Al2O3 [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 (FexAu1−x), from cyanide tailings. The gold and iron in iron powders are separated by sulfuric acid leaching. Then the leachate is used to produce FeSO4·7H2O 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 Table1. 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) (Figure1) 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, 8, x FOR PEER REVIEW 3 of 14 Components TFe Au * SiO2 Al2O3 CaO MgO Na2OK2O Components TFe Au* SiO2 Al2O3 CaO MgO Na2O K2O Content (%) 44.96 3.83 24.35 2.47 0.58 0.67 1.26 1.15 ComponentsContent (%) 44.96TFe 3.83Au* 24.35SiO2 Al2.472O 3 0.58CaO MgO0.67 Na1.262O 1.15K2O Content (%) 44.96 3.83* The 24.35 unit of * AuThe2.47 was unit g/t. of0.58 Au was0.67 g/t. 1.26 1.15 * The unit of Au was g/t. 1 1- hematite 1 2-1- quartzhematite 1 2- quartz 1 2

2 1 1 Intensity (a.u.) Intensity 1 1 11 1 1 Intensity (a.u.) Intensity 1 2 1 1 11 1 2 2 2 1 1 1 2 1 2 2 1 1 10 20 30 40 50 602 70 801 1 90

10 20 30 40 2θ50 (°) 60 70 80 90 Figure 1. XRD pattern of2θ received (°) cyanide tailings. 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 2.2. DirectThe recoveryReduction process Roasting was of Cyanideshown inTailings Figure 2. Cyanide tailings were dried at 105 °C overnight. ◦ The30 g recovery tailings were process thoroughly was shown mixed in with Figure coal2 . and Cyanide CaO with tailings a designated were dried dosage. at 105 Herein,C overnight. the The recovery process was shown in Figure 2. Cyanide tailings were dried at 105 °C overnight. 30 g tailingsdosage of were coal thoroughly or CaO was mixed expressed with by coal a andpercentage CaO with that a referred designated to its dosage. mass ratio Herein, to cyanide the dosage 30 g tailings were thoroughly mixed with coal and CaO with a designated dosage. Herein, the of coaltailings. or CaO The wasdosage expresseds of coal and by CaO a percentage were chosen that in the referred range from to its 8.0 mass wt.% ratioto 20 wt. to% cyanide and from tailings. 5 dosage of coal or CaO was expressed by a percentage that referred to its mass ratio to cyanide The dosageswt.% to 20 of wt. coal%,and respectively. CaO were The chosen mixture in was the rangeadded fromto a graphite 8.0 wt.% crucible to 20 wt.%with a and lip. fromThe direct 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 was conducted in a muffle furnace (CD-1400X, Zhengzhou Chida Tungsten & 20 wt.%,wt.% respectively.to 20 wt.%, respectively. The mixture The was mixture added was to aadded graphite to a cruciblegraphite crucible with a lip. with The a lip. direct The reductiondirect roastingMolybdenum was conducted Products in Co. a mufﬂe Ltd., Zhengzhou, furnace (CD-1400X, China). When Zhengzhou the reduction Chida temperature Tungsten &reached Molybdenum its reductiondesignated roasting value of was 1050 conducted‒1250 °C , thein acrucible muffle wasfurnace transferred (CD-1400X into, theZhengzhou furnace and Chida maintained Tungsten for & Products Co. Ltd., Zhengzhou, China). When the reduction temperature reached its designated value Molybdenum90 min at a reduction Products atmosphere Co. Ltd., Zhengzhou,. After that, Chinathe crucible). When was the taken reduction out of temperaturethe furnace and reached cooled its of 1050–1250 ◦C, the crucible was transferred into the furnace and maintained for 90 min at a reduction designateddown to room value temperature of 1050‒1250 in air °C. , the crucible was transferred into the furnace and maintained for atmosphere.90 min atAfter a reduction that, the atmosphere crucible was. After taken that, out th ofe crucible the furnace was andtaken cooled out of down the furnace to room and temperature cooled in air.down to room temperature in air. cyanide tailings 3.83 100 Gold grade /g/t Yield /% 44.96 TFe content /% cyanide taCiloinagl s+ C3a.O83 100 Gold grade /g/t Yield /% 44.96 TFe content /% direct reduction Coal + CaO mdaigrneect icre sdeupcatiroantion

8.23 45.34 0.19 51.33 irmona gpnoewtdice srsepatraailtiinogns 92.43 5.94 8.23 45.34 0.19 51.33 iroanc pido wledaechrsingtailings 92.43 5.94 216.58 1.72 28.71 leaching resaidcuide leachlienagchate

216.58 1.72 cyanidation leaching residue 28.71 extraction crysletalclihzatieon

7.10 1.72 cyanidation / cyan eidxetracptiorengnant crystallization / 222.42 residue solution FeSO4·7H2O 19.49 7.10 1.72 / pregnant / 222.42 Figure 2. Flowsheet for the recovercyanidye of gold and iron from cyanide tailings at the reduction residue solution FeSO4·7H2O 19.49 temperature of 1200 °C . (The Yield is defined as the weight percentage of solid products to received cyanide tailings as shown in Equation (5)). FigureFig 2.ureFlowsheet 2. Flowsheet for thefor recovery the recover of goldy of andgold iron and from iron cyanide from cyanide tailings tailingsat the reduction 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 Figure2 were collected as magnetic separation tailings. The TFe or gold recovery of reduced iron powders was calculated as follows: W × β ε = 1 1 × 100% (1) W0 × β0 where, ε was the TFe or gold recovery (%), W1 was the weight of reduced iron powders (g), β1 was the TFe content (%) or gold grade (g/t) of iron powders, W0 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 H2SO4 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 FeSO4·7H2O. 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.

W − W η = 1 3 × 100% (2) W1

W × β − W × β γ = 1 1 3 3 × 100% (3) W1 × β1 where, η and γ were the weight loss ratio (%) and iron leaching rate (%), respectively, W3 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: W × β − W × β φ = 3 3 5 5 × 100% (4) W3 × β3 where, φ was the gold extraction rate (%), β3 and β5 was the gold grade (g/t) of acid-leaching residue and cyanide residue, respectively, W5 was the weight (g) of cyanide residue. In this process, the yield (δi) of solid products to cyanide tailings in Figure1 was defined as Equation (5): Wi δi = × 100% (5) W0 where, Wi (i.e., W0, W1, W2, W3, W4 and W5) was the weight (g) of cyanide tailings, iron powders, magnetic separation tailings, acid-leaching residue, FeSO4·7H2O crystals and cyanide residue, respectively. Metals 2018, 8, 561 5 of 13 Metals 2018, 8, x FOR PEER REVIEW 5 of 14

2.2.5.5. Analysis Analysis and and Characterization Characterization The Au Au grade grade of of solid solid products products was determinedwas determined by inductively by inductively coupled coupled plasma-optical plasma emission-optical emissionspectrometry spectrometry (ICP-OES, (ICP iCAP-OES, 7000, iCAP Thermo 7000, Thermo Fisher ScientificFisher Scientific Inc., Waltham, Inc., Waltham, MA, USA) MA, USA at radio) at rfrequencyadio frequency (RF) power(RF) power of 1.35 of 1.35 kW, kW, air flowair flow rate rate of 14of 14 mL/min, mL/min, and and argon argon auxiliary auxiliary flowflow raterate of 0.3 mL/min. 0.3 g samples were digested with 3 mL ionized water, 9 mL HCl and 3 mL HNO3 at 130 0.3 mL/min. 0.3 g samples were digested with 3 mL ionized water, 9 mL HCl and 3 mL HNO3 °atC 130for 15◦C min for in 15 a min closed in amicrowave closed microwave acid digest acider ( digesterDS-360x, (DS-360x, Grand Analytical Grand Analytical Instrument Instrument Co. Ltd., Guangzhou,Co. Ltd., Guangzhou, China). The China). TFe content The TFe of contentsolid products of solid w productsas analyzed was analyzedby a titanium by a titanium(III) chlorid (III)e reductionchloride reduction and potassium and potassium dichromate dichromate titration titration method method (Chinese (Chinese standard, standard, GB/T GB/T6730.65 6730.65-2009).-2009). The contents of SiO2, Al2O3, CaO, MgO, Na2O and K2O in cyanide tailings, and of Mn, Pb, As, Si, Al, and The contents of SiO2, Al2O3, CaO, MgO, Na2O and K2O in cyanide tailings, and of Mn, Pb, As, Si, Cu in FeSO4·7H2O crystals were also obtained by ICP-OES. The mineral compositions of cyanide Al, and Cu in FeSO4·7H2O crystals were also obtained by ICP-OES. The mineral compositions of tailingscyanide, roasted tailings, products roasted products and acid and-leaching acid-leaching residue were residue investigated were investigated by XRD by(DMax XRD-RB (DMax-RB,, Rigaku Corporation,Rigaku Corporation, Tokyo, JapanTokyo,) under Japan) the under conditions the conditions of 40 kV, of 40 100 kV, mA 100 and mA Cu and Kα Cu radiation. Kα radiation. The morphologyThe morphology of roasted of roasted product productss was was performed performed with with scanning scanning electron electron microscopy microscopy (SEM (SEM,, EVO18 EVO18,, Carl Zeiss Group Corporation, Oberkochen, Germany).Germany). Roasted product productss w wereere mounted on epoxy resinresin and further polished before the SEM observation.

3. Results Results and and Discussion Discussion

3.13.1.. Recovery Recovery of of Gold Gold and and Iron Iron with Direct Reduction Roasting

3.1.13.1.1.. Effect Effect of of Reduction Temperature As shown in in Fig Figureure 3 3,, thethe TFeTFe contentcontent andand recoveryrecovery ofof reducedreduced ironiron powderspowders werewere increasedincreased fromfrom 68.84% 68.84% to to 92.43% 92.43% and and from from 78.98% 78.98% to to 93.19% 93.19%,, respectively respectively,, as the temperature was raised from from ◦ 1050 to to 1200 1200 C.°C. It It suggests suggests that that high high temperature temperature can promotecan promot thee reduction the reduction of iron of oxides iron to oxides metallic to metalliciron. Meanwhile, iron. Meanwhile, the gold recoverythe gold was recovery increased was from increased 74.33% from to 95.03%, 74.33% indicating to 95.03%, the indicating enrichment the of ◦ enrichmentremained gold of remained in the tailings gold into in the iron tailings powders. into As iron the powders. temperature As the was temperature raised up to was 1250 raisedC, the up gold to 1250recovery °C, was the furthergold recovery increased was to 95.70%. further It increased clearly indicated to 95.70%. that higherIt clearly temperature indicated could that promote higher temperaturethe capture of could more promote gold by the in situ capture generated of more metallic gold by iron. in situ generated metallic iron.

100 100

90 95 90 80 85 70 80 TFe content (%) content TFe TFe content

TFe recovery (%) of metal Recovery 60 75 Gold recovery 50 70 1050 1100 1150 1200 1250 Reduction temperature (°C)

FigFigureure 3. EffectEffect 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 min,, coal dosage of 20 wt. wt.%% and CaO dosage of 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 Fe2O3→Fe3O4 →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 O can be further reduced to FeO. As the reduction of FeO to metallic iron is a strong endothermic3 reaction,4 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 SiO in cyanide tailings, FeO can react with SiO to form fayalite, a low melting melting point material promot2 ing to form molten slag [36]. The amount2 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

Metals 2018, 8, x FOR PEER REVIEW 6 of 14 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 Figure3, 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 revealed3.1.2. Effect that of higher Coal Dosage coal dosage could promote to recover gold and iron from cyanide tailings. The high coal dosage can increase the C/O (fixed carbon/oxygen) ratio in the reduction system, whichAs enhance showns in the Figure reduction4, when atmosphere the coal dosage and promotes was increased the reduction from 8 wt.% of more to 20 FeO wt.%, to metallic the TFe andiron [3gold9]. Thus, recovery the washigher rapidly amount increased of metallic from iron 56.36% can to surely 93.03% capture and from more 42.53% gold.to The 93.19%, FeO in respectively. reduction systemHowever, is thea substance TFe content decreasing of iron powders the liquidus was just temperature increased fromby forming 89.92% a to fayalite 92.43%.-type When slag the [3 dosage8,40]. Whenwas further more raisedFeO was up toreduced 25 wt.%, to theFe at TFe higher content C/O slightly ratio, decreased,the molten butslag almost became no less change, which of the might TFe reduceand gold the recovery mechanical was mixing observed. of gold Thus, and the slag optimum in slag coal phase dosage. Therefore, was 20 wt.%. the allocation The result of revealed gold in metallicthat higher iron coal phase dosage should could be enhanced. promote to recover gold and iron from cyanide tailings.

100 100

90 95 80

70 90 60

TFe content (%) content TFe TFe content 50 85

TFe recovery Recovery of metal (%) Gold recovery 40 80 30 8 12 16 20 24 28 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, whichAs enhances shown in the Fig reductionure 5, when atmosphere the CaO and dosage promotes was the increased reduction from of more 5 wt. FeO% to to 20 metallic wt.%, ironthe [TFe39]. recoveryThus, the was higher raised amount slightly, of metallic but the iron Au can recovery surely capturesignificantly more gold.was raised The FeO from in reduction77.65% to system 97.46%. is Thea substance TFe content decreasing showed the a week liquidus relationship temperature with by CaO forming dosage. a fayalite-type Thus, the optimum slag [38 CaO,40]. Whendosage more was 20FeO wt. was%. The reduced result to revealed Fe at higher that the C/O CaO ratio, dosage the had molten a remarkable slag became effect less, on which the gold might recovery. reduce the mechanical mixing of gold and slag in slag phase. Therefore, the allocation of gold in metallic iron 100 100 phase should be enhanced.

3.1.3. Effect of CaO Dosage 95 95 As shown in Figure5, 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 (%) content TFe 85 TFe content TFe recovery 80 (%) of metal Recovery Gold recovery 80 75 4 8 12 16 20 CaO dosage (wt.%)

Metals 2018, 8, x FOR PEER REVIEW 6 of 14 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 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. However, the TFe content of iron powders was just increased from 89.92% to 92.43%. When the dosage was further raised up to 25 wt.%, the TFe content slightly decreased, but almost no change of the TFe and gold recovery was observed. Thus, the optimum coal dosage was 20 wt.%. The result revealed that higher coal dosage could promote to recover gold and iron from cyanide tailings. The high coal dosage can increase the C/O (fixed carbon/oxygen) ratio in the reduction system, which enhances the reduction atmosphere and promotes the reduction of more FeO to metallic iron [39]. Thus, the higher amount of metallic iron can surely capture more gold. The FeO in reduction system is a substance decreasing the liquidus temperature by forming a fayalite-type slag [38,40]. When more FeO was reduced to Fe at higher C/O ratio, the molten slag became less, which might reduce the mechanical mixing of gold and slag in slag phase. Therefore, the allocation of gold in metallic iron phase should be enhanced.

100 100

90 95 80

70 90 60

TFe content (%) content TFe TFe content 50 85

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%. TheMetals TFe2018 content, 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.

100 100

95 95

90 90

85 TFe content (%) content TFe 85 TFe content TFe recovery 80 (%) of metal Recovery Gold recovery 80 75 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 Table1. 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.

θ SiO2 + 2FeO = Fe2SiO4, ∆GT = −33231 + 12.26 T (6)

θ Fe2SiO4 + 2C = 2Fe + SiO2 + CO, ∆GT = 167700 − 160.91 T (7) θ Fe2SiO4 + 2C + CaO = 2Fe + CaO·SiO2 + CO, ∆GT = 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 Figure6 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 Figure6, 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 8 of 14 Weight percent of Fe 1550 ℃ ) 1538 1550 L 0 5 Weight10 15 percent20 30 of Fe40 50 607080100 ℃ ℃ ) 1500 1538 1600 L

0 5 10 15 20 30 40 50 607080100 ℃ 1538℃ 1500 1600 1450 (δ-Fe)

1433℃ 97

14331538℃ ℃ (δ-Fe) 98 1450 (δ-Fe)

1400 L 89 97 1394℃ 1433℃ 97

1433℃ (δ-Fe) 1400 L+(γ-Fe) 98

1394℃ ℃ Temperature ( (γ-Fe) 1400 L L+(γ-Fe)89 97 1394 1400 L+(γ-Fe) 1350 1394℃ Temperature ( (γ-Fe) 1200 57 L+(γ-Fe)1173℃ 94 96 98 100 (γ-Fe) 1350 74 92 Atomic percent of Fe (%) 1173℃ 1200 1064.43℃ 57 94 96 98 100 (γ-Fe) 74 92 950 Atomic percent of Fe (%)

1064.43℃ ) 1000

912℃ ℃ 18.5% 1038℃ 98 950 (γ-Fe) ℃ 912 ) 1000 868℃ 900

Temperature (°C)

912℃ ℃ 18.5% 1038℃ 53 96 (γ-Fe) 98 912℃ (Au)+(α-Fe)868℃ 99 900 Temperature (°C) 800 769℃ 99 (Au) ℃ 770℃ 850 868 53 96 (Au)+(α-Fe)(Tc) (α-Fe)99 (α-Fe) 800 769℃ 99 (Au) 850 868℃ 770℃ (Au)+(α-Fe)

(Tc) (α-Fe) 800 (α-Fe) Temperature ( 600 c 770℃ (Au)+(α-Fe)769℃ (T ) 0 10 20 30 40 50 60 70 80 90 100 800 Temperature ( 750 600 c ℃ 769℃ (T ) 770 0 Au10 20 30Atomic40 50percent60 70of Fe80 90 100Fe 97 98 99 100 750 Atomic percent of Fe (%) Au Atomic percent of Fe Fe 97 98 99 100 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 ◦ wasFigureFig generatedure7 7 showed showed even XRDXRD at 1050 patternspatterns °C . The ofof roastedroasted intensity productsproducts of diffraction obtainedobtained peaks fromfrom assigned 10501050 toto 1250 1250to metallic °CC.. Metallic iron became iron ◦ wasstrong, generated revealing even higher at 1050 amount °CC.. The and intensity larger ofcrystal diffraction size of peaks metallic assigned iron at to elevated metallic temperature. iron became As strong,shown revealing revealing in Figure higher higher 8, the amount amountgrain size and and of larger largermetallic crystal crystal iron size (grayish size of ofmetallic metallicwhite) iron at iron 1050at elevated at °C elevated was temperature. just temperature. about 20 As µm , ◦ µ Asshownand shown mostin Fig in of Figureure these 8, 8the ,grains the grain grain were size size separated.of of metallic metallic Nevertheless,iron iron (grayish (grayish white)some white) grain at at 1050 1050 size °C ofC was wasmetallic just just about aboutiron at 20 20 1200 µmm, , °C ◦ andwas most increased of these up grains to about were 100 separated.separated. µm and Nevertheless, dispersed metallic some irongrain particles size of metallic had been iron aggregated at 1200 °C Cinto µ waslarger increased grain s. up Herein, to about it should 100 µmm point and dispersedout that metallic metallic iron iron grains particles containing hadhad been gold aggregated is recovered into by largerthe grinding grains.grains. Herein, and magnetic itit shouldshould separation point point out out thatprocess. that metallic metallic The ironlarger iron grains grainsgrain containing sizecontaining means gold easygold is recovered separationis recovered by of theby iron grindingthegrains grinding andfrom and magnetic the magnetic slags separation by theseparation phys process.ical process. separation The largerThe methodlarger grain grain sizesuch means sizeas magnetic means easy separationeasy separation. separation of ironTherefore, of grains iron as fromgrainsshown the from slags in Figthe by ureslags the 3, physicaltheby thehigher phys separation Auical and separation Fe method recovery suchmethod of as iron magneticsuch powders as magnetic separation. was observed. separation. Therefore, Therefore, as shown as in Figureshown3 in, the Fig higherure 3, the Au higher and Fe Au recovery and Fe of recovery iron powders of iron was powders observed. was observed. 1 2 1-metallic iron 1 2-pseudowollastonite 2 1-metallic iron 2 2-pseudowollastonite3-wollastonite 2 4-fayalite 2 2 3-wollastonite 2 1 2 2 4-fayalite 2 1 1250C 2 1 2 22 1 1250C 2 32 3 2 2 4 4 23 1 1200C 1 2 2 4 43 32 1 4 3 223 1 1200C 4 3 3 2 3 1150C 1 Intensity (a.u.) 4 2 3 3 3 1 33 2 4 3 34 2 33 3 1150C 1 Intensity (a.u.) 1 3 3 3 3 3 1 3 4 4 3 3 3 3 3 3 1 1100C 3 3 1 4 4 3 3 3 3 3 3 1 4 1100C 1 3 3 4 3 3 4 3 3 3 3 3 1 1050C 4 1 3 3 4 3 3 3 3 3 1 1050C 20 30 40 50 60 70 80 90 20 30 40 50 60 70 80 90 2θ (º) 2θ (º) Figure 7. XRD patterns of roasted products obtained at different roasting temperatures. Figure 7. XRDXRD patterns of roasted product productss obtained at different roasting temperatures.temperatures.

Metals 2018, 8, 561 9 of 13 Metals 2018, 8, x FOR PEER REVIEW 9 of 14 Metals 2018, 8, x FOR PEER REVIEW 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 of roasted products obtained at 1050°C◦ (a) and 1200°C◦ (b) with same Figuremagnification 8. SEM of 200 images. of roasted products obtained at 1050 C (a) and 1200 C (b) with same magnificationmagniﬁcation of 200 200.. 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 shown, had in Figure lower7 ,fraction as the roastingin slag phase temperature and disappeared was increased, at 1250 the °C . fayalite, At the same a low time, melting more point material, had lower fraction in slag phase and disappeared at 1250 °C . At◦ the same time, more pointwollastonite material, (α- hadCaSiO lower3) in fraction slag phase in slag was phase transformed and disappeared to pseudowollastonite at 1250 C. At the (β-CaSiO same time,3) at wollastonite (α-CaSiO3) in slag phase was transformed to pseudowollastonite (β-CaSiO3) at moretemperature wollastonite higher ( αthan-CaSiO 11503) °C in slagsince phase the transformation was transformed temperature to pseudowollastonite of α-CaSiO3 to (ββ--CaSiOCaSiO3 3was) at temperature higher than 1150 °C since the transformation temperature of α-CaSiO3 to β-CaSiO3 was temperature1126 °C . Compared higher than to fayalite 1150 ◦C ( sincemelting the pointtransformation of 1205 °C temperature), (pseudo)wollastonite of α-CaSiO tohadβ-CaSiO the highwaser 1126 °C . Compared to fayalite (melting point of 1205 °C ), (pseudo)wollastonite3 had the high3 er 1126melting◦C. point Compared of 1540 to fayalite°C [35]. (meltingThus, the point amount of 1205 of liquidus◦C), (pseudo)wollastonite in reduction system had would the higher remarkable melting melting point of 1540 °C [35]. Thus, the amount of liquidus in reduction system would remarkable pointdecrease of 1540 as more◦C[35 pseudowollastonite]. Thus, the amount was of liquidus generated in reduction at higher system reduction would temperature, remarkable decreasepreventing as decrease as more pseudowollastonite was generated at higher reduction temperature, preventing morethe mechanical pseudowollastonite mixing of was molten generated gold inat higher slag phase. reduction Thus, temperature, more gold capreventingn be allocated the mechanical into iron the mechanical mixing of molten gold in slag phase. Thus, more gold can be allocated into iron mixingphase to of molten form Au gold-Fe in alloy. slag phase. It was Thus, well more in agreement gold can be with allocated higher into Au iron recovery phase to format elevated Au-Fe phase to form Au-Fe alloy. It was well in agreement with higher Au recovery at elevated alloy.temperature It was well(Figure in agreement 3) and with with increased higher CaO Au recovery dosage (Fig at elevatedure 5). temperature (Figure3) and with temperature (Figure 3) and with increased CaO dosage (Figure 5). increased CaO dosage (Figure5). 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 temperatureSulfuric on acid iron leaching dissolution was used was toshown separate in Fig goldure and 9. As iron the in temperature iron powders. was The raised effect from of leaching 30 to 50 temperature on iron dissolution was shown in Figure 9. As the temperature was raised from 30 to 50 temperature°C, the weight on loss iron ratio dissolution and iron was leaching shown ratein Figure were9 .increased As the temperature from 71.41% was to raised 96.20% from and 30from to °C, the weight loss ratio and iron leaching rate were increased from 71.41% to 96.20% and from 5082.97%◦C, the to 98.82%, weight lossrespectively. ratio and W ironhile leaching the temperature rate were was increased further fromraised 71.41% to 90 °C to, 96.20%the iron and leaching from 82.97% to 98.82%, respectively. While the temperature was further raised to 90 °C , the iron leaching 82.97%rate became to 98.82%, almost respectively. constant. AtWhile acid the-leaching temperature temperature was further of 50 raised °C , the to 90 yield◦C, the of acid iron- leaching rate became almost constant. At acid-leaching temperature of 50 °C , the yield of acid-leaching rateresidue became to cyanide almost tailings constant. was At acid-leachingreduced to 1.72%, temperature and the of gold 50 ◦ gradeC, the reached yield of acid-leachingas high as 216.58 residue g/t residue to cyanide tailings was reduced to 1.72%, and the gold grade reached as high as 216.58 g/t to(Fig cyanideure 1), tailingsrevealing was remarkable reduced to gold 1.72%, enrichment and the gold and grade significant reached weight as high decrease as 216.58 of g/tacid (Figure-leaching1), (Figure 1), revealing remarkable gold enrichment and significant weight decrease of acid-leaching revealingresidue. remarkable gold enrichment and signiﬁcant weight decrease of acid-leaching residue. residue.

100 100 100 100

90 90 90 90

80 80 80 80 Weight loss ratio Weight loss ratio

70 70 Weight loss ratioWeight (%) loss

70 Iron leaching rate 70 Iron leaching rate (%) Weight loss ratioWeight (%) loss Iron leaching rate Iron leaching rate (%)

60 60 60 30 40 50 60 70 80 90 60 30 40 50 60 70 80 90 Leaching temperature (°C) Leaching temperature (°C) Figure 9. Effect of acid-leaching temperature on iron dissolution of reduced iron powders. FigFigureure 9 9.. EEffectffect of of acid acid-leaching-leaching temperature temperature on iron dissolution 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

MetalsFigure 2018, 8, x10 FOR showed PEER REVIEW XRD pattern of acid-leaching residue. The diffraction peaks could be assigned10 of 14 to cohenite (Fe3C) and bustamite (CaMn(SiO3)2), generated in the direct reduction roasting. A broad Apeak broad with peak a 2θ withdegree a 2θ of degree 10–40 ◦ ofC was10–40 observed, °C was observed, indicating indicating the appearance the appearance of amorphous of amorphous materials. materials.However, However, no diffraction no diffraction peak of metallic peak of iron metallic was observed, iron was revealing observed nearly, revealing complete nearly dissolution complete ◦ dissolutionof metallic ironof metallic in dilute iron H 2inSO dilute4 solution H2SO at4 solution 50 C. Therefore, at 50 °C . Therefore, it can reasonably it can reasonably infer that theinfer Au-Fe that thealloy Au structure-Fe alloy in structure iron powders in iron must powders be completely must be completely broken, and broken, gold should and gold be fully should exposed be fully in exposedacid-leaching in acid residue.-leaching In this resid work,ue. In metallic this work, iron inm ironetallic powders iron in wasiron found powders to readily was found react withto readily dilute ◦ reactH2SO with4 to producedilute H FeSO2SO4 to4 at produce 50 C according FeSO4 at 50 to Equations°C according (9–11). to Eq However,uations (9 sulfuric–11). However, acid leaching sulfuric of acidiron oxidesleaching (e.g., of iron hematite) oxides should(e.g., hematite) be conducted should at be much conducted higher temperatureat much higher (90-110 temperature◦C) and with(90‒ 110longer°C ) timeand with (2-4 h)longer [44,45 time]. (2‒4 h) [44,45]. Fe(s) + H2SO4(l) = FeSO4(l) + H2(g) (9) Fe(s) + H2SO4(l) = FeSO4(l) + H2(g) (9) 4FeSO4(l) + 2H2SO4(l) + O2(g) = 2Fe2(SO4)3(l) + 2H2O(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)

2 1 1-Cohenite 2 2 2-Bustamite 2 1 111 1 1 1

2 1 Intensity (a.u.) Intensity

10 20 30 40 50 60 70 80

2θ ()

FigFigureure 10. XRD pattern of acid acid-leaching-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 FeSO4·7H2O 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 FeSO ·7H O to cyanide tailings was as high as 222.42% as shown in Figure 1. The chemical composition4 in2 Table 2 showed that the content of 222.42% as shown in Figure1. The chemical composition in Table2 showed that the content of FeSO4·7H2O reached 96.73% with low content of harmful elements such as Pb and As. The obtained FeSO4·7H2O reached 96.73% with low content of harmful elements such as Pb and As. The obtained FeSO4·7H2O crystals can be used as raw materials in the production of pigment, water treatment FeSO ·7H O crystals can be used as raw materials in the production of pigment, water treatment chemicals4 2and 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 FeSO4·7H2O crystals. Components FeSO4·7H2O Mn Pb As Si Al Cu ContentComponents (%) FeSO96.734·7H 2O0.011 Mn 0.002 Pb 0.0001 As Si0.03 0.01 Al 0.003 Cu Content (%) 96.73 0.011 0.002 0.0001 0.03 0.01 0.003 3.4. Recovery of Gold from Acid-Leaching Residue 3.4. RecoveryThe enriched of Gold gold from in Acid-Leaching acid-leaching Residue residue was finally extracted by the cyanidation extraction. As givenThe enrichedin Table 3 gold, the ingold acid-leaching extraction rate residue was wasincreased ﬁnally at extracted higher cyanide by the dosage cyanidation. At the extraction. cyanide dosageAs given of in 40 Table kg/t,3 ,the the extraction gold extraction rate was rate increased was increased up to at96.72% higher, which cyanide revealed dosage. nearly At the complete cyanide exposuredosage of of 40 gold kg/t, in the acid extraction-leaching rateresidue was. Th increasede gold in up pregnant to 96.72%, solution which was revealed not further nearly recovered complete atexposure the current of gold stage, in acid-leaching and further studies residue. such The as gold adsorption in pregnant by solutionactivated was carbon not furtheror ion exchange recovered of at goldthe current would stage, be tak anden. furtherAs shown studies in Figure such as 1, adsorptionthe gold grade by activated of produced carbon cyanid or ione exchangeresidue reached of gold would7.10 g/t be, which taken. could As shown be recycled in Figure into1, the the gold direct grade reduction of produced roasting cyanide as raw residue materials. reached Therefore, 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 (Table3)[ 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 FeSO4·7H2O 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

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