Production of Direct Reduction Grade Iron Concentrate from Pickling Sludge by Reduction Calcination and Magnetic Separation

Abstract

Taking pickling sludge generated from the steel rolling process at a Chinese steel mill as the subject of study, this research investigates a combined mineral processing and metallurgical process involving gasification-reduction roasting, magnetic separation, and waste acid recovery. The main elements in this pickling sludge are Fe and Cl, with contents of 46.50% and 12.70%, respectively. The primary component is chlorine-containing iron oxide, and a significant amount of amorphous material is also present. The study investigated the effects of various calcination temperatures, calcination times, reducing agent dosages, and material thicknesses on gasification-reduction performance indicators. The results indicate that using a reduction calcination–grinding–magnetic separation process, with coal as the reducing agent, a calcination temperature of 1100 °C, a reducing agent dosage of 15%, and a calcination time of 2 h, the chlorine volatilization rate exceeds 97%. Furthermore, when the roasted ore is ground to a particle size where 85% passes through a −0.076 mm screen and is recovered via magnetic separation, an iron concentrate with a grade of over 69.50% can be obtained. This iron concentrate meets the quality requirements for high-grade iron concentrates used in direct reduced iron (DRI) production. The small amount of tailings from the iron concentration process can be utilized in the production of bricks, cement, and other products, thereby achieving the efficient comprehensive utilization of acid washing sludge.

1. Introduction

China is the world’s largest steel producer, with an annual crude steel output exceeding 1 billion tons. The steel production process inevitably generates large quantities of byproducts, such as blast furnace slag, blast furnace dust, steel slag, steelmaking dust, and acid pickling sludge from rolling mills. The annual volume of these solid wastes exceeds 100 million tons [1,2,3]. In addition to being classified as general industrial solid waste, this sludge and dust also contains a significant amount of hazardous waste. If not properly disposed of or utilized, it will not only occupy large tracts of land for storage, posing long-term environmental risks, but also result in a massive waste of resources. Therefore, the steel industry should actively explore technologies for the recycling and reuse of solid waste to build new types of steel enterprises that are green, low-carbon, energy-efficient, emission-reducing, technologically advanced, and environmentally friendly. In parallel with waste valorization efforts, fundamental process innovations are also being pursued to reduce the carbon footprint of steel production. For instance, Taimullah et al. recently demonstrated the feasibility of producing AISI 300 series stainless steel directly from lateritic nickel and chromite ores via hydrogen plasma smelting reduction, achieving complete metallization within 120 s without fossil-based carbon reductants [4]. Such advances underscore the global trend toward sustainable and low-carbon metallurgical technologies.

During the cold rolling process in steel production, the use of hydrochloric acid pickling to remove oxide scale from the surface of steel generates large amounts of waste acid containing Fe²⁺, Fe³⁺, and excess HCl. When treating this type of waste acid through spray calcination or acid regeneration processes, neutralization, oxidation flocculation, and filter press dewatering result in the generation of large amounts of pickling sludge. This pickling sludge contains approximately 45% iron and 15% chlorine and is classified as hazardous waste. Tang et al. investigated the use of acid pickling filter cakes generated from steel mills, which were neutralized and mixed with alkaline dust collection ash to serve as sintering feedstock [5,6]. Tang et al. directly added stainless steel pickling sludge to oxide pellets for resource recovery [7]. Meng et al. conducted research on the use of pickling sludge, processed through drying, briquetting, and drying, as a slag-forming agent in electric arc furnace steelmaking [8]. Traditionally, both domestically and internationally, the primary methods for treating this type of acid leaching filter cake have included direct landfilling, chemical neutralization, and calcination [9,10,11,12,13,14,15,16,17]. Direct landfilling not only wastes a significant amount of iron resources but also contaminates soil and groundwater due to the high chlorine content in the filter cake. Although chemical neutralization can reduce the acidity of acid-washed sludge, it also generates new pollutants during the process, which still require further treatment. If the sludge is fed directly into the sintering process, the chlorine present will corrode the equipment and reduce the strength of the sintered ore, thereby affecting the quality of subsequent blast furnace ironmaking [18,19]. The calcination method is an effective process for treating this type of hazardous waste. By adding a reducing agent to promote the transformation of the iron and chlorine phases, it facilitates solid–solid and solid–gas phase transformations and separations. At an appropriate calcination temperature, iron is converted into a solid form—either as magnetite or metallic iron—within the calcined ore, while chlorine is reduced to hydrogen chloride or chlorine gas and extracted in the gas phase. This gas is then absorbed by an alkaline solution to prevent pollution. The kinetics of gas–solid reduction reactions are strongly dependent on temperature and the reducing atmosphere [20,21]. Zhang et al. employed a high-temperature reduction–magnetic separation process to treat stainless steel pickling sludge. At 1350 °C, they obtained magnetic products containing Fe, Cr, and Ni, with an Fe recovery rate exceeding 90% and reducing the heavy metal Cr and Ni content in the tailings to levels compliant with discharge standards [22]. Lin et al. studied pyrometallurgical treatment processes for pickling sludge from eight companies and proposed environmental protection measures, including exhaust gas treatment and safety regulations [23]. Noboru et al. used graphite as a reducing agent and employed microwave heating for the direct reduction of pickling sludge, also yielding an Fe-Cr-Ni alloy [24].

However, the products obtained in the aforementioned studies have a complex elemental composition and low iron grade, making it impossible to achieve high-value-added utilization. Therefore, this study addresses the challenge of disposing of hazardous pickling sludge generated during the steel rolling process in the steel industry. Based on an analysis of the solid waste’s characteristics, the study proposes the use of reduction roasting, magnetic separation, waste acid recovery processes to produce direct-reduction-grade iron concentrate. This approach not only enables the high-value-added resource utilization of iron from the pickling sludge but also comprehensively addresses environmental considerations such as exhaust gas treatment and regenerated acid recovery, thereby providing a reference framework for the recycling of similar hazardous wastes.

2. Results and Discussion

2.1. Gasification Reduction Calcination

2.1.1. Calcination Temperature

To investigate the effects of varying calcination temperatures on iron reduction indices and chlorine volatilization, gasification-reduction roasting tests were conducted at different temperatures. The reduction agent dosage was 15%, the calcination time at 2 h, and the material thickness at 25 mm. The calcined ore was water-quenched, ground to a particle size distribution where 85% of the particles were smaller than 0.076 mm, and then subjected to magnetic separation. The results of calcination temperature tests are presented in Figure 1. The results demonstrate that with increasing temperature, both the Fe grade and recovery rate of the magnetic concentrate increased, accompanied by a gradual increase in the Cl volatilization rate, indicating more complete reduction of iron oxides and enhanced chlorine removal. At a reduction temperature of 900 °C, the iron concentrate grade was 61.42%, the iron recovery rate was 86.53%, and the chlorine volatilization rate was 81.51%. When the temperature was raised to 1100 °C, the chlorine volatilization rate exceeded 97%, indicating effective removal of chlorine. At this point, the iron grade of the concentrate was 69.52%, meeting the quality standards for direct reduction grade iron concentrate, with an iron recovery rate of 96.75%, representing relatively ideal iron reduction and recovery performance. Further increasing the temperature promotes chlorine volatilization and increases the iron concentrate grade, though the improvement is not significant. This indicates that higher temperatures accelerate the transformation of mineral phases, causing chlorine-bearing minerals to transition from the solid phase to the gaseous phase more rapidly. Simultaneously, high temperatures promote the conversion of the iron oxide phase and ferric chloride phase into the magnetite phase, while also increasing the grain size of the magnetite crystals, which is beneficial for improving iron grade and recovery rate. Taking into account both product indicators and energy consumption, the roasting and reduction temperature was set at 1100 °C.

2.1.2. Calcination Time

A sufficient amount of time is required for a chemical reaction to proceed fully. To investigate the effects of varying calcination times on iron reduction indices and chlorine volatilization, experiments were conducted with different calcination times during gasification-reduction. The reducing agent dosage was 15%, the calcination temperature was 1100 °C, and the material thickness was 25 mm. The calcined ore was water-quenched, ground to a particle size distribution where 85% of the particles were smaller than 0.076 mm, and then subjected to magnetic separation. The test results are shown in Figure 2. The results show that as the calcination time increases, the Fe grade and recovery rate of the iron concentrate tend to rise, while the volatilization rate of Cl in the volatiles also increases. At a calcination time of 1 h, the iron concentrate grade was 61.99%, the iron recovery rate was 83.83%, and the chlorine volatilization rate was 81.81%. This indicates that under relatively short roasting times, iron reduction was insufficient, chlorine volatilization was poor, and phase transformation reactions were incomplete. When the roasting time was extended to 2 h, the iron grade of the iron concentrate reached 69.52%, the iron recovery rate was 96.75%, and the chlorine volatilization rate exceeded 97%. Further increases in roasting time resulted in only minor improvements in chlorine volatilization and iron concentrate grade. After comprehensive consideration, the roasting time was set at 2 h.

2.1.3. Reductant Agent

The reduction of iron oxides and chlorides requires the addition of sufficient reducing agents and the attainment of specific temperature conditions. To investigate the effects of varying reducing agent dosages on iron reduction rates and chlorine volatilization during calcination, experiments were conducted using gasification reduction calcination with different reducing agent (pulverized coal) dosages. The roasting time was 2 h, the roasting temperature was 1100 °C, and the material thickness was 25 mm. The ore was water-quenched, ground to a particle size distribution where 85% of the particles were ≤0.076 mm, and then subjected to magnetic separation. The experimental results are shown in Figure 3. The results show that when the reducing agent dosage was 5%, the Cl volatilization rate was relatively low at 91.25%, the Fe grade of the iron concentrate was 63.85%, and the Fe recovery rate was 84.13%. This indicates that the reducing agent dosage was insufficient to meet the requirements of the reduction reaction. As the amount of reducing agent added increases, both the Cl volatilization rate and the Fe grade of the iron concentrate rise. When the reducing agent dosage is 15%, the Cl volatilization rate exceeds 97%, and the Fe grade of the iron concentrate exceeds 69%. Further increases in the reducing agent dosage do not result in significant changes in the Cl volatilization rate, nor do they lead to significant improvements in the Fe grade or recovery rate of the iron concentrate. Taking all factors into consideration, the optimal reducing agent dosage is determined to be 15%.

2.1.4. Material Thickness

To investigate the effects of different material thicknesses on iron reduction indices and chlorine volatilization, roasting tests were conducted with varying feed bed thicknesses. To prepare samples with different material thicknesses, a specified mass of the homogenized mixture (pickling sludge + 15% pulverized coal) was weighed and uniformly loaded into a corundum crucible. The mixture was gently tapped to ensure a flat and even surface without compaction. The material thickness was controlled by adjusting the mass of the mixture loaded. The roasting time was 2 h, the roasting temperature was 1100 °C, and the reducing agent dosage was 15%. The roasted ore was water-quenched, ground to a particle size distribution where 85% of the particles were ≤0.076 mm, and then subjected to magnetic separation. The test results are shown in Figure 4. The results show that as the material thickness increases, the Fe grade and recovery rate of the iron concentrate tend to rise, while the chlorine volatilization rate gradually decreases. This indicates that the thickness of the calcination layer influences iron reduction and chlorine volatilization to a certain extent. When the material thickness was low, the good permeability facilitated chlorine volatilization; however, it also caused the reducing gas from the reducing agent to volatilize too rapidly, shortening the contact time with the high-valent iron requiring reduction and thereby affecting iron reduction. As the material thickness increases, iron reduction becomes more complete, and both the iron concentrate grade and recovery rate increase; however, excessively high material thickness impedes chlorine volatilization, resulting in excessive chlorine content in the iron concentrate. Considering the overall effects of iron reduction and chlorine volatilization, the roasting material thickness was selected as 25 mm.

2.1.5. Cooling Method for Roasted Ore

Two primary methods are used to cool products after reduction roasting. One is water quenching cooling, in which the high-temperature roasted ore is immersed directly in water. The other is furnace cooling, in which heating is stopped once the roasting time is complete, and the roasted ore is left in the tube furnace to cool naturally to room temperature. Two cooling methods were tested to investigate their impact on the performance indicators of magnetic reduction calcination. The calcination tests were conducted under the following conditions: a reduction time of 2 h, a reduction temperature of 1100 °C, a coal content of 15%, and a material thickness of 25 mm. Magnetic separation tests were conducted on the roasted ore under conditions where 85% of the ground ore had a particle size of −0.076 mm, with the test results shown in

Table 1. Effect of cooling method on test index.

Cooling Method	Product	Yield (%)	Grade (%)		Fe Distribution Rate (%)	Cl Volatilization Rate (%)
			Fe	Cl
Furnace cooling	Iron concentrate	64.87	68.98	0.37	96.27	1.88
	Tailings	13.52	12.82	0.51	3.73	0.54
	Volatile matter	21.61	/	57.61	/	97.58
	Feed	100.00	46.48	12.76	100.00	100.00
Water quenching cooling	Iron concentrate	65.11	69.52	0.38	96.75	1.93
	Tailings	13.45	11.30	0.52	3.25	0.55
	Volatile matter	21.44	/	58.23	/	97.52
	Feed	100.00	46.78	12.80	100.00	100.00

. The results show that the water-quenching cooling method yields higher iron concentrate grades and recovery rates, with an iron grade of over 69.5% and a recovery rate of 96.75%, demonstrating clear advantages in the recovery and utilization of iron resources. This indicates that water serves to isolate the ore from air, preventing the roasted ore from being re-oxidized, and that the water-quenching cooling method was both rapid and efficient as well as easy to operate [25]. The chlorine volatilization rates for both cooling methods were similar, both exceeding 97.5%, indicating that there was little difference between them. Taking all factors into consideration, the water quenching cooling method was deemed more suitable.

2.2. Roasted Ore Grinding Magnetic Separation Test

The key factor affecting mineral separation is the degree of liberation among the minerals, which is achieved through the grinding process. Therefore, the roasted ore obtained through water quenching cooling was ground to different fineness degrees and subjected to magnetic separation tests using Davis tube, with a magnetic field strength of 100 mT as described above. The test results are shown in Figure 5. The results show that as the grinding fineness of the roasted ore increases, the Fe grade of the iron concentrate gradually rises, while the Fe recovery rate decreases accordingly. At a grinding fineness where 65% of the particles were −0.076 mm, the Fe grade of the iron concentrate was only 66.49%, indicating that the grinding particle size was relatively coarse and the degree of mineral liberation was poor. When the grinding fineness reaches 85% at −0.076 mm, an iron concentrate with a grade of 69.61% can be obtained. Further grinding did not significantly improve the iron grade of the concentrate.

2.3. Study of Calcination Reaction Mechanism

To elucidate the calcination reaction mechanism, X-ray diffraction analysis (Rigaku D/Max—1200 X-ray diffractometer, Tokyo, Japan) was performed on the calcined ore; the results are shown in Figure 6. Additionally, scanning electron microscopy and energy dispersive spectroscopy analyses (SEM-EDS, FEI Company, Hillsboro, OR, USA) were conducted on both the raw ore and the calcined ore; the results are shown in Figure 7. The results indicate that, following calcination of the acid washing sludge, most of the iron oxide underwent a reduction reaction, resulting in the formation of magnetite. Combined with the calcination test results, the chlorine volatilization rate reached over 97%, indicating that the chlorine in the acid washing sludge underwent chemical reactions and was released as gas. The primary chemical reaction equations that may have occurred are shown in Equations (1)–(8).

FeCl₃·6H₂O → FeCl₃ + 6H₂O

(1)

FeCl₃ + 3H₂O → Fe(OH)₃ + 3HCl ↑

(2)

FeCl₂ + 2H₂O → Fe(OH)₂ + 2HCl ↑

(3)

4Fe(OH)₂ + O₂ + 2H₂O → 4Fe(OH)₃

(4)

2Fe(OH)₃ → Fe₂O₃ + 3H₂O

(5)

2C + O₂ → 2CO

(6)

C + CO₂ → 2CO

(7)

3Fe₂O₃ + CO → 2Fe₃O₄ +CO₂

(8)

3. Materials and Methods

3.1. Materials

The acid pickling sludge samples were obtained from a steel company in Anhui Province, China, and were generated during the cold-rolled steel acid pickling and waste acid regeneration processes, with an annual production volume exceeding 3000 tons. Due to its strong acidity and corrosiveness, it is classified as a difficult-to-treat solid waste. The main components of the pickling sludge were determined by chemical titration, with the results shown in

Table 2. Main components of pickling sludge (wt.%).

Component	Fe	SiO2	Al2O3	CaO	MgO	Cl	P	S
Content	46.50	1.27	0.64	0.04	0.03	12.70	0.04	0.04

. The results indicate that the primary valuable elements in the acid pickling sludge are iron and chlorine, and it consists of a mixture of compounds such as FeCl₃, FeCl₂, Fe(OH)₃, and Fe(OH)₂. X-ray diffraction (Rigaku D/Max—1200 X-ray diffractometer, Tokyo, Japan) analysis of the pickling sludge was performed. The test conditions were as follows: copper target, power of 50 kV and 150 mA, step length of 0.02°, and a scanning range of 5–80°. The analysis results are shown in Figure 8. The results indicate that its main components were Al₂O₃, Fe₈O₈(OH)₈Cl_1.35, and Zn(SO₄)(H₂O)₇. The results of the thermogravimetric analysis of the acid-washed sludge are shown in Figure 9. The results indicate that the sample weight decreases as the temperature rises. No weight loss occurred between 0 and 50 °C. Significant weight loss was observed between 50 and 450 °C. Above 450 °C, the rate of weight loss slowed, and by 1300 °C, the decrease was minimal.

Coal was used as the reducing agent. The proximate analysis of the coal was carried out according to the Chinese national standards GB/T 212-2008 and GB/T 214-2007 [26,27]. The chemical composition of the coal ash was determined by X-ray fluorescence spectrometry (Bruker S8 TIGER, Karlsruhe, Germany) and chemical titration methods. The results of the coal’s industrial analysis and the chemical composition analysis of its ash are shown in

Table 3. Industrial analysis and chemical composition results of anthracite coal (wt.%).

Fixed Carbon	Ash Content	Volatile Matter	Moisture Content	S	P
80.22	11.56	7.78	0.25	0.01	0.002

and

Table 4. Analysis results of the main chemical components of coal ash (wt.%).

Component	SiO2	Al2O3	CaO	MgO	Fe
Content	43.13	19.55	18.46	1.31	4.42

, respectively. The results show that the coke contains 80.22% fixed carbon, 11.56% ash, and 7.78% volatile matter, with low S and P content. The ash in the coal consists primarily of SiO₂, Al₂O₃, CaO, and a small amount of iron, indicating that this coal is a suitable calcined reducing agent.

3.2. Methods

Weigh a specific amount of pickling sludge and combine it with a corresponding proportion of reducing agent (pulverized coal), with the pulverized coal having a particle size of −1 mm. After thorough mixing using a dry mixing method, place the mixture in an alumina crucible. To promote chlorine volatilization and maintain a reducing atmosphere, the crucible is left open, and the mixture is calcined in a tube furnace equipped with a vacuum pump. Vacuum extraction was initiated simultaneously, with the extracted gas collected in a flask containing an alkaline solution. After a specified calcination time, the crucible was removed from the furnace, and the calcined sand was rapidly quenched with water to cool it. The cooled roasted ore is ground in an XZM-100 mill (Wuhan exploration machinery, Wuhan, China) until 85% of the particles are below 0.076 mm. Iron concentrate is then obtained by separation using an XCGS-Φ50 Davis tube (Wuhan exploration machinery, Wuhan, China) with a magnetic field strength of 100 mT. A schematic diagram of the roasting experimental setup is shown in Figure 10, and the experimental process flow is shown in Figure 11.

4. Conclusions

(1)

The primary valuable elements in the pickling sludge are iron and chlorine. The experimental results demonstrate that the pickling sludge can be efficiently recycled and utilized through the reduction roasting–magnetic separation process. During the reduction roasting process, calcination temperature, reducing agent dosage, and calcination time significantly influence the reduction performance, while the grinding fineness of the roasted ore is also a critical factor affecting the grade of the iron concentrate.

(2)

Using a reduction calcination–grinding–magnetic separation–spent acid recovery process, with a calcination temperature of 1100 °C, a reducing agent dosage of 15%, and a calcination time of 2 h, the chlorine volatilization rate reaches over 97%. When the roasted ore is ground to a particle size distribution where 85% of the particles are −0.076 mm, magnetic separation yields iron concentrate with a grade of over 69.50%, meeting the quality requirements for direct reduction-grade iron concentrate.

(3)

The key to gasification-reduction calcination lies in maintaining a stable reaction atmosphere within the furnace. This process must ensure that Fe₂O₃ is reduced to magnetite and retained in the slag phase to produce high-quality iron concentrate, while also taking into account the volatilization and recovery of chlorine. The iron tailings obtained after magnetic separation can be utilized in building materials such as cement and bricks, thereby achieving the safe treatment and disposal of pickling sludge.