Novel Process for Comprehensive Utilization of Iron Concentrate Recovered from Zinc Kiln Slag

Abstract

A novel process, which includes hydrochloric acid leaching, iron powder cementation, sulphide precipitation and spray pyrolysis, was proposed to treat iron concentrate recovered from zinc kiln slag for comprehensive utilization. Treated by two-stage hydrochloric acid leaching, the leaching extents of Ag, Pb, Cu, Fe and Zn are 96.78%, 95.14%, 97.72%, 94.51% and 87.74%, respectively. More than 99% Cu and Ag are recovered from the leach liquor when iron powder is 25% higher than the stoichiometric requirement. More than 96% Pb and Zn is removed with three times of theoretical consumption of FeS, and the concentration of impurities of the final solution is less than 500 mg/L. Fe₂O₃ powder, which is spherical with a mean size of 12 μm and purity of 99%, is prepared with final solution at 700 °C. In this process, not only the metal values can be recovered effectively, but also the iron resources of zinc kiln slag can be converted into high purity Fe₂O₃, thus, the comprehensive utilization of iron concentrate recovered from zinc kiln slag can be realized.

Introduction

The output of zinc was 6.15 million tonnes in China in 2015, which accounted for approximately 43% of global zinc output [1]. And in China, about 85% zinc was produced by zinc hydrometallurgy [2]. Zinc hydrometallurgy has developed rapidly in China, almost all zinc hydrometallurgy technologies have been applied in Chinese zinc smelting plants. These technologies have their own advantages and disadvantages [3]. Until now, zinc hydrometallurgy mainly consists of the traditional roast-leach-purifying-electrowinning process. Only a few enterprises leach zinc sulfide concentrate directly to produce zinc. In the leaching process, zinc oxide and zinc sulfate in calcine are dissolved into solution, the leaching rate of zinc could reach 80–85%. And almost all Pb, Au, Ag, In, Ge, ZnFe₂O₄ and 60% Cu enter the leaching residue [4]. Consequently, hydrometallurgical zinc residues not only contain 18–26% zinc, but also contain valuable metals such as Pb, Ge, In. To improve the effect of comprehensive utilization of resources, zinc residues will be treated by a fuming method through vaporization in rotary kiln (also called the Waelz process) [5], in which coke serves as the fuel and reductant. Pb, Cd, In, Ge, Ga in zinc residues volatilize into the gas phase at temperature of 1100–1300 °C and enrich in zinc oxide powders for recovery subsequently. Meanwhile, involatile matter such as Fe, Cu, SiO₂, are left in the kiln slag.

At present, physical separation is the main method for handling kiln slag. Typical physical separation methods include crushing-ball milling-magnetic and gravity separation [4], or flotation-magnetic separation-flotation [6]. Among these physical separation methods, the most representative and widely used one is “separation of carbon by flotation-separation of iron by magnetic separation” [7–11]. Coke, iron concentrate and tailings can be obtained using this process. When the iron content in the iron concentrate recovered from zinc kiln slag reaches the minimum standard (T_Fe ≥ 55%) of “GB/T 25953-2010 Recovery of iron concentrates by beneficiation of nonferrous metals”, enterprises usually sell the qualified iron concentrate to the iron works for the use of ore proportioning [12], or sell the concentrate to lead smelters for replacing pyrite cinder, or it can be used as sulfur-capturing agent in reduction-matting smelting [13]. Among these applications, the concentrate is widely used as an iron-making raw material. Because FeS, CuS and other sulphur-containing compounds in the kiln slag are tightly intermingled with unreacted Fe₃O₄. The iron concentrate separated by magnetic separation contains high content of sulphur, and the sulphur content is well above quality standards of iron concentrate for using in iron-making raw material, which requires the content of impurity sulphur to be lower than 0.8%. The high sulphur-containing concentrate could make steels hot short and decrease in ductility and toughness [14]. Besides, the energy used by desulphurization will increase due to high content of sulphur in raw materials. Experience shows that coke ratio would increase 5% when the sulphur content in raw materials increases 1% [15]. The content of arsenic is about 0.4–1.5% in the iron concentrate recovered from zinc kiln slag, which also restricts the concentrate’s application in iron and steel industry [16]. Silver occurs mostly as silver sulfide with a fine particle size in zinc kiln slag, and the silver sulfide is usually encapsulated in FeS and Fe₃O₄. Besides, a small amount of silver can substitute metal atoms of metallic iron and other sulfide phases to form isomorphous compounds [17]. Silver mainly deports to the iron concentrate separated by physical separation. There is no detrimental effects for ironmaking process with silver-containing iron concentrates. But it will waste the precious metal resources, if the silver is not recovered.

Based on the present situation, a new process for the treatment of the iron concentrate was developed, which consists of the following steps: hydrochloric acid leaching, recovering Cu and Ag by iron powder cementation, recovering Pb and Zn by sulfide precipitation, and preparing iron powders by spray pyrolysis. In the process, the effective recoveries of Ag, Cu, Zn, Pb and the high value utilization of iron from the iron concentrate can be realized.

Experiment

Experimental Material

The iron concentrate used in the present study was obtained from Zinc Smelting Company of Shangluo in Shanxi Province, China. The results of chemical and phase analyses are shown in Table 1 and Fig. 1. Table 1 indicates that the main constituent of the iron concentrate is Fe, and its content is up to 56.4%. The XRD pattern of the kiln slag iron concentrate sample, as presented in Fig. 1, shows only four main phases of Fe₃O₄, CuFe₂O₄, Fe₂O₃ and SiO₂ with no evidence of other phases, probably due to their less content or lower crystallinity. The results of wet screening, that the particle size of the iron concentrate is fine and 80% of the particle sizes are less than 200 mesh, indicate that the iron concentrate can be leached directly without further grinding.

Table 1

The main chemical compositions of iron concentrate used in the present study (wt%)

Fe	Zn	Cu	Pb	S	SiO₂	Al₂O₃	MnO	Ag
56.4	2.51	2.51	0.99	4.75	4.61	1.85	1.28	265 g/t

../images/468727_1_En_146_Chapter/468727_1_En_146_Fig1_HTML.gif — Fig. 1
The XRD pattern of iron concentrate used in the present study

Experimental Process and Principle

Experimental process of this study is shown in Fig. 2. The leaching agent in this study is hydrochloric acid solution, which provides Cl⁻ to coordinate with Cu⁺/Cu²⁺, Ag⁺, Pb²⁺, Zn²⁺ and thus valuable metals can be leached from the iron concentrate. Then Cu and Ag, enriched in the leaching solution, are displaced by iron powder to obtain copper and silver-bearing residues. Subsequently, the solution is deeply purified by FeS to obtain lead and zinc-bearing residues. Finally, the solution could be used as starting material for spray pyrolysis process to produce iron oxide and hydrochloric acid recovered from tail gas in the process can return to the leaching process.

../images/468727_1_En_146_Chapter/468727_1_En_146_Fig2_HTML.gif — Fig. 2
The process scheme for comprehensive utilization of iron concentrate

Experimental Method

Hydrochloric Acid Leaching

One liter of certain concentration of hydrochloride acid solution was put into a three neck flask (2 L). Experimental apparatus were assembled properly and then the hydrochloride acid solution was heated and agitated. When the reaction temperature reached 60 °C the iron concentrate (100 g) was added into the three neck flask and the leaching reaction began. In the leaching process, the reaction solution was sampled at 30 min intervals and its chemical compositions were analyzed. The total leaching time was 120 min. When the leaching experiment was finished, the solid and liquid in the reaction system were separated and analyzed.

Leaching Solution Treatment

The leaching solution (0.5 L) was put into a three neck flask (1 L), and then the solution was heated and agitated. When the reaction temperature reached 30 °C a certain amount of iron powder was added into the flask. The residue and solution in this system were separated by filtration and analyzed after a certain reaction time. Finally, the solution, obtained from iron powder cementation process, was put into a three neck flask again and reacted with FeS at target temperature. The sulfide residues and solution in this system were separated by filtration and analyzed after a certain reaction time.

Spray Pyrolysis

The main equipment of spray pyrolysis experiment was self-made ultrasonic atomization spray pyrolysis furnace. The experimental methods were listed as follows:

First, the furnace was heated to 500–900 °C, and carrier air was introduced to the furnace for 30 min to ensure temperature uniformity. Then the solution, obtained from sulfide precipitation, was atomized by ultrasonic. The atomized droplets, carried by air, went into the furnace and started the thermal decomposition reaction. Finally, the reaction products were collected and analyzed. In this study, the atomizing power and flow rate of carrier air were 2.0 W and 125 L/h respectively.

Sample Characterization

The total iron and ferrous iron contents were determined by potassium dichromate titration. ICP-AES was applied to determine the content of Ag, Cu, Pb, Zn, Si in the leaching solution and purified solution. The iron concentrate, leaching residue, and spray pyrolysis product were characterized via powder XRD analysis (RIGAKU-TTRIII, Cu/Kα, λ = 1.5406Å). Morphological studies of the synthesized products were carried out directly by a JEOL jsm-6360 scanning electron microscope.

Results and Discussion

Tow-Stage Countercurrent Leaching

The iron concentrate was treated by two-stage countercurrent leaching on the basis of prophase experimental results. The low-acid leaching experimental conditions were listed as follows: leaching time of 120 min, HCl concentration of 3 mol/L, reaction temperature of 60 °C, and liquid to solid ratio of 10:1 mL/g. The following conditions were about high-acid leaching experiment: leaching time of 120 min, HCl concentration of 6 mol/L, reaction temperature of 60 °C, liquid to solid ratio of 5:1 mL/g. The results of two-stage countercurrent leaching are shown in Table 2.

Table 2

The leaching rates in two-stage countercurrent leaching process

Elements	Fe	Ag	Cu	As	Pb	Zn
Low-acid leaching rates (%)	60.38	79.22	80.39	49.61	83.81	57.56
High-acid leaching rates (%)	34.13	17.56	17.33	46.19	11.33	30.18
Total leaching rates (%)	94.51	96.78	97.72	95.80	95.14	87.74
Final concentration (g/L)	121.17	0.067	6.26	1.12	1.92	5.69

Table 2 indicates the total leaching rates are high, and the leaching rates of Ag, Cu, As, Pb, Zn, Fe were 96.78%, 97.72%, 95.80%, 95.14%, 87.74% and 94.15% respectively. The iron concentrate can reach 121.17 g/L in low-acid leaching solution, and Ag, Cu, Zn are enriched in the solution to a high extent. The recoveries of valuable metals are realized. Meanwhile, most of As and Pb went into the leaching solution. Table 3 shows that the acid concentration is as low as 0.19 mol/L in the final leaching solution, which provides convenience for further purification of the leaching solution and enables the acid in the leaching solution to get efficient use.

Table 3

The final acid concentration in two-stage countercurrent leaching process

Acid leaching types	High-acid leaching	Low-acid leaching
Final acid concentration (mol/L)	3.57	0.19

The XRD patterns of the leaching residues of low-acid and high-acid leaching process are shown in Fig. 3. As is seen in Fig. 3a, the XRD pattern of low-acid leaching residues shows diffraction peaks of SiO₂, FeSiO₃, FeS, ZnS, Fe₃O₄, which indicates FeSiO₃, FeS, ZnS, Fe₃O₄ cannot be leached completely. And SiO₂ is the principle component of quartz in the raw materials, which are unable to be leached. The XRD pattern of high-acid leaching residues, as presented in Fig. 3b, shows only three phases of SiO₂, FeS, ZnS, which indicates FeSiO₃, Fe₃O₄ can be leached completely by high-acid leaching. The leaching rates of Fe, Zn are relatively low, because there are still some parts of FeS, ZnS cannot be leached by high-acid leaching.

../images/468727_1_En_146_Chapter/468727_1_En_146_Fig3_HTML.gif — Fig. 3
The XRD patterns of low-acid leaching residues (a) and high-acid leaching residues (b)

Separation and Extraction of Valuable Components

Iron Powder Cementation

On the basis of standard reduction potential, Cu²⁺, Ag⁺, Pb²⁺, H⁺ except for Zn²⁺ can react with iron.

Hydrogen evolution is not obvious due to high pH value (around 1.5) in the leaching solution. The thermodynamic calculation results of the main iron powder cementation reactions, presented in Fig. 4, show the reaction sequence of the main ions is Ag⁺>Cu²⁺>H⁺>Pb²⁺ and the reaction tendency with iron of Ag⁺, Cu²⁺ is higher. In theory, copper and silver could be enriched by iron powder cementation method with controlled amount of iron. The effects of reaction time on the cementation efficiency of Ag, Cu, Pb were investigated under the condition that reaction temperature was 30 °C, stirring speed was 300 r/min, initial pH value of the solution was 1.5, excess coefficient of iron powder was 1.25. The result is shown in Fig. 5.

../images/468727_1_En_146_Chapter/468727_1_En_146_Fig4_HTML.gif — Fig. 4
Thermodynamics calculation results of the main displacement reactions

../images/468727_1_En_146_Chapter/468727_1_En_146_Fig5_HTML.gif — Fig. 5
The effect of reaction time on the recovery rates of Cu, Ag and Pb

Figure 5 indicates the cementation rates of Ag and Cu could reach 99.97% and 99.62% respectively, when the reaction time is 20 min. At 20–40 min, there is little influence on the cementation rates of Cu, Ag with the increase of reaction time. But the cementation rate of Cu would decrease if the reaction time is over 45 min, probably due to the re-dissolution of copper. And iron can also replace lead in the solution, but the cementation rate (around 5%) is quite low. Considering the recovery efficiency of Ag, Cu, the reaction time should be set as about 30 min.

Recovery of Lead by FeS

Almost all Cu, Ag and a little Pb in the solution can be removed by iron powder cementation, but it has little removal effect on Zn. The solution obtained from the above process can be deeply purified by sulfide precipitation, based on the small solubility products of sulfides. The logC-pH relationships of the main sulfides in the reaction system are shown in Fig. 6. The stabilization region of FeS, shown in Fig. 6, is far less than ZnS, PbS, Ag₂S and CuS in acidic solution. When the pH of the solution is lower than −0.18, the precipitation sequence of the sulfides in the reaction system is Ag₂S > CuS > PbS > ZnS > FeS. When −0.18 < pH < 4.38, the sulfides precipitated in an order of CuS > Ag₂S > PbS > ZnS > FeS. And the precipitation sequence is CuS > Ag₂S > ZnS > PbS > FeS, if pH > 4.38. Namely, Cu, Ag, Pb, and Zn can be deeply removed by adding FeS in acidic condition. The effects of FeS on the removal of Cu, Ag, Pb and Zn in the solution obtained from iron powder cementation are investigated, based on the as-mentioned principle. The results are shown in Fig. 7.

../images/468727_1_En_146_Chapter/468727_1_En_146_Fig6_HTML.gif — Fig. 6
Log C–pH relationship in M_xS_y-H₂O system(298.15 K)

../images/468727_1_En_146_Chapter/468727_1_En_146_Fig7_HTML.gif — Fig. 7
The effect of stoichiometric ratio of FeS on the removal rate of Cu, Ag, Pb and Zn

Figure 7 indicates the removal effects of Cu, Ag, Pb, Zn are distinguished, and the removal order is Ag, Cu, Pb, Zn, which is consistent with theoretical calculations in Fig. 6. The removal rates of Cu and Ag can reach above 96%, with 1.5 times theoretical amounts of FeS, and the concentrations of Cu, Ag in the purified solution are lower than 1 ppm. And 97.33% Pb and 96.00% Zn can be removed, when 3 times theoretical amounts of FeS are added into the solution. Finally, impurities in the purified solution are less than 500 mg/L, which lays a good foundation for preparing subsequent iron oxide product.

Spray Pyrolysis

High purity FeCl₂ solution could be obtained through iron powder cementation and FeS purification of the initial leaching solution. High purity iron oxide could be prepared by spray pyrolysis method using the FeCl₂ solution, and the hydrochloric acid recovered from tail gas in spray pyrolysis process can return to the leaching process for reuse. The main reaction in the process is:

$4 {\text{FeCl}}_{2} + {\text{O}}_{2} + 4 {\text{H}}_{2} {\text{O}}\mathop \rightarrow \limits^{\Delta } 2{\text{Fe}}_{2} {\text{O}}_{3} {\text{ + 8HCl}}{ \uparrow }$

(1)

The SEM image and XRD pattern of Fe₂O₃ product obtained at 700 °C are shown in Fig. 8. As seen in Fig. 8, near-spherical Fe₂O₃ powder with D₅₀ = 12 μm could be produced. The chemical composition analysis results of the as-synthesized Fe₂O₃ sample are listed in Table 4. The impurity contents, listed in Table 4, are less than 0.5% in the Fe₂O₃ obtained at 700 °C, which can satisfy the production requirements of anticorrosive paints and magnetic materials.

../images/468727_1_En_146_Chapter/468727_1_En_146_Fig8_HTML.gif — Fig. 8
The SEM image and XRD pattern of Fe₂O₃ obtained at 700 °C

Table 4

The XRF results of Fe₂O₃ product

Elements	Zn	Pb	Si	Mn	Ca	Mg	Al	O	Fe
Content (%)	0.029	0.012	0.027	0.063	0.0166	0.0024	0.0067	31.16	68.68

Conclusions

(1)
The leaching effect of the iron concentrate recovered from zinc kiln slag is good in hydrochloric solution, due to the coordination ability of chlorine ion. The leaching retes of Ag, Pb, Cu, Fe and Zn can reach 96.78%, 95.14%, 97.72%, 94.51% and 87.74% respectively by use of two-stage countercurrent leaching method.
(2)
The Cu and Ag in the leaching solution can be effectively enriched through iron powder cementation, but the remove effects are bad for Pb and Zn. The recovery rates of Cu and Ag could reach above 99%, with one point twenty-five times theoretical amount of iron powder in the iron powder cementation unit.
(3)
The solution, obtained from iron powder cementation, can be deeply purified by FeS. More than 96% of Pb and Zn could be recovered in the sulfide precipitation process, when three times theoretical amount of FeS were added into the solution. Finally, impurities in the purified solution were less than 500 mg/L.
(4)
When the spray pyrolysis temperature is 700 °C, near-spherical Fe₂O₃ powder with D₅₀ = 12 μm and purity > 99% could be produced.

Acknowledgements

This work was supported by the National Natural Science Foundation of China [grant numbers: 51404307].

References

1.
The depth report in zinc industry. Fortune Securities, 2016-06-07
2.
Li RG (2010) Status and development of lead and zinc smelting process in China. China Nonferrous Metall 16:13–20. https://doi.org/10.3969/j.issn.1672-6103.2010.06.003Crossref
3.
Peng RQ (2005) Zinc metallurgy. Central south University Press, Chang Sha
4.
Wang H (2007) Intergraded cycling utilization of volatilizing kiln slag during zinc hydrometallurgy. China Nonferrous Metall 6:46–50. https://doi.org/10.3969/j.issn.1672-6103.2007.06.014Crossref
5.
Matthes J, Waibel P, Keller HB (2011) A new infrared camera-based technology for the optimization of the waelz process for zinc recycling. Miner Eng 24(8):944–949. https://doi.org/10.1016/j.mineng.2011.04.020Crossref
6.
Xie DY (2004) Experimental study on recovering silver from zinc volatilizing kiln slag. MS thesis, Central South University
7.
Li J, Niu H, Peng JH (2008) Present situation and prospect about comprehensive utilization of zinc kiln slags. Multip util Miner Resour 6:44–48. https://doi.org/10.3969/j.issn.1000-6532.2008.06.013Crossref
8.
Fu ZJ (1988) Discussion on treatment method of volatilization kiln slag in zinc hydrometallurgy. Nonferrous Met 1:41–46
9.
Wang H (2007) Study of recovery of value metals from waelz clinker by physical separation. Rare Met Cem Carbides 3(1):31–34. https://doi.org/10.3969/j.issn.1004-0536.2007.01.008Crossref
10.
Ma J, Jiao HG, Pan LY (2009) Research on recycle and reuse zinc volatile kiln slag by magnetic separation. J Henan Polytech Univ (Nat Sci) 28(6):788–791. https://doi.org/10.3969/j.issn.1673-9787.2009.06.020
11.
Schriefer J (1997) Reaping the value from dust and slag. Iron Age New Steel 13(2):24
12.
He SC, Liu ZH, Liu ZY (2013) Thermodynamic analysis of leaching of iron concentrate separated from kiln slag in zinc hydrometallurgy. Chin J Nonferrous Met 23(12):3430–3439Crossref
13.
Feng ZW, Lu AM (2009) The research and practice of technology about comprehensive recovering coal, silver, iron from zinc volatile kiln slag. Resour Recycl 9:56–57
14.
Huang XH (1981) Metallurgy principle of iron and steel. Metallurgical Industry Press, Bei Jing
15.
Zhu MY (2005) Modern metallurgy. Metallurgical Industry Press, Bei Jing
16.
Chen GL (2014) Novel process for dearsenization, desulfurization, desilverization in iron concentrate recovered from zinc kiln slag. MS thesis, Kunming University of Science and Technology
17.
Liu X (2015) Experimental research on recovery of C and Ag from zinc kiln slag by flotation. Hunan Nonferrous Met 31(1):16–19. https://doi.org/10.3969/j.issn.1003-5540.2015.01.005Crossref

Novel Process for Comprehensive Utilization of Iron Concentrate Recovered from Zinc Kiln Slag

Abstract

Keywords

Introduction

Experiment

Experimental Material

Experimental Process and Principle

Experimental Method

Hydrochloric Acid Leaching

Leaching Solution Treatment

Spray Pyrolysis

Sample Characterization

Results and Discussion

Tow-Stage Countercurrent Leaching

Separation and Extraction of Valuable Components

Iron Powder Cementation

Recovery of Lead by FeS

Spray Pyrolysis

Conclusions

Acknowledgements