With the rapid development of infrastructure projects and economic growth, China has become one of the largest steel producers in the world. This has led to increasing demand for raw iron ores over the recent years, however, there are not enough high-grade iron ores due to over-exploitation. A low-grade mineral deposit with huge reserves, e.g., oolitic iron ores [1, 2], has not been effectively utilized as a potential resource. It would be desirable to utilize the low-grade iron ore by applying effective processing treatments (e.g., roasting [3, 4], flotation ), to overcome the shortage  and to cover the demand of the steel industry.
However, for the beneficiation of typical oolitic iron ore, the concentrating process is challenging, because of its low grade Fe, poor liberation of iron minerals  and complex mineralogical composition [8, 9, 10]. Commercially, satisfactory mineral processing that can effectively utilize the ore is stilf under way due to industrial cost. Generally, direct beneficiation of low-grade iron ore without pretreatment is not appropriate, as the iron oxide mineral (e.g., Fe2O3) in the iron ore has weak magnetic properties.
It is possible and usually necessary to improve the material properties before the beneficiation process by pretreatment [11, 12], using various roasting techniques including microwave roasting , alkali roasting , chlorination roasting  and so on. For iron ore and other secondary resources  or solid waste [17, 18], the processes of roasting  and magnetic separation  have sometimes been proved to be effective. These studies investigated the various conditions and the influencing factors. However, the microscopic mechanism and how these components behave were not fully interpreted in a quantitative way.
The present study was undertaken to gain a better understanding of the changing behaviors of the iron phase and reduction area with roasting time during the magnetic process, by quantitatively examining the microstructures and growth behavior. The effects of roasting temperature, reduction time, particle size, and coal type have also been investigated in detail. The present study aims to provide a basis for the commercial utilization of such refractory iron ores, for the technological development and industrial process control.
2 Materials and methods
2.1 Raw materials and chemicals
The raw ores were oolitic iron from west Hubei province, China, taken by the notch sampling method, with size ≤ 150 mm. Then the ores were crushed to be less than 6 mm, by a Jaw Crusher through an open circuit.
After that, samples with size ≤2.0 mm were obtained by a closed-loop circuit using a Roller Grinding machine. The chemical components of the sample are listed in Table 1, TFe (Total Fe content) is ~49%, and Ig (Mass loss of ignition) is 6.5%.
WG coal powder with size ≤1.0 mm was applied as a reducing agent in the magnetic roasting process, to change the iron phase, and therefore to modify the magnetic properties for the facilitation of magnetic separation . The main industrial values of WG coal are shown in Table 2, from which it can be seen that the fixed carbon content of this coal is high as ~78%, while the volatile matter and moisture are low, making it a suitable candidate as reductant.
2.2 Analytical methods and parameters
For the study of the magnetic roasting process and its influencing factors a muffle furnace (KSY–12D–16) was used, with the roasting results checked and validated by a magnetic tube (XCGS-73). Please note that the magnetic tube is widely used in laboratory scale experiments to determine the content of magnetic components for strong magnetic ores. For large-scale separation, in order to mimic real industrial practice, a drum type weak magnetic separator (Φ400×400) was applied, which can handle a large quantity of feeding.
The detailed procedure was as follows: the oolitic iron sample was thoroughly mixed with the reducing agent (WG coal powder), then the mixture was put into a high temperature heat-resistant stainless container, which was then roasted using a muffle furnace at selected temperatures. After a certain time of roasting, the product was processed by water quenching, dewatering, drying and grinding, finally its components were magnetically separated using a magnetic tube.
Observation and a counting method were conducted to examine the statistical ratio of the reduction area for the roasted oolitic iron, under various roasting conditions with different sizes, using a polarizing microscope (Leiz, Germany). The roasted iron sample was first added to special cement (as an adhesive medium), then was cut into test block of 40 mm×25 mm, by rough grinding, fine grinding and polishing, finally it was investigated under microscope.
Fe recovery (ε) is the percentage of the useful component within the separated concentrate to that of the original oolitic iron sample, which can be calculated as, ε = γ × β/α, where γ is production yield, β is Fe content of the product and α is Fe content of the original oolitic iron.
A conical ball mill, XMQ-Φ150×50 was applied for lab scale grinding during the magnetic tube test, and XMQ-Φ240×90 for the case of the drum type weak magnetic separator. X-ray diffraction (XRD) was performed using Rigaku D/max-3c (Japan), voltage 40 kV, current 50 mA, scanning speed of 2 (degrees)/min.
The conducted research is not related to either human or animals use.
3 Results and discussion
3.1 Mineralogical characteristics and mineral compositions
The raw iron ores are mainly in oolitic shape from the microscopic view (Figure 1), the oolitic particles were composed of hematite and clay minerals by layered structure, with a gangue showing a circling layer distribution (Figure 1a). Some Silicon-quartz may exist inside the oolitic particle (Figure 1b), after roasting, the oolitic iron particle (hematite) can be transformed into magnetite.
The inner structure of the oolitic particles was dependant on the particle size, most were 3 ~ 5 µm dense concentric wrapped shells, with siliceous quartz or clay like Kaolinite in the middle (Figure 1b). In the raw iron sample, the Al minerals mainly include Chlorite, Kaolinite, and the inclusion of Collophanite. The Mg minerals are mainly Dolomite, Chlorite and Calcite.
The XRD pattern of the raw oolitic iron ore is shown in Figure 2, it was seen that the main minerals are hematite, ankerite, chlorite, collophanite, a small amount of Kaolinite and Quartz. These are in agreement with the observation from the microscope. The mineral components are listed in Table 3, with the content of the three main minerals being 64%, 13% and 10%, respectively.
The Fe mineral phase and its distribution is shown in Table 4, the main constitute is hematite-limonite which account for 91% of total Fe. The recoverable Fe in the oolitic iron ore is mainly in the form of magnetite and hematite-limonite, theoretically the Fe recovery (ε) can be 91.23%. Moreover, some part of Fe in Siderite can be recovered; however, the iron within the Ferric silicate cannot be utilized.
3.2 Overall results of magnetic roasting for oolitic iron ore
Usually, the roasting temperature and roasting time are the most important factors for the reduction reaction in the process of magnetic roasting, i.e. the chemical reaction would lead to the phase transformation of Fe mineral phases and cause the change of magnetic property.
To study the effect of roasting temperature, the fixed parameters were as follows: the particle size was less than 2 mm, the sample amount was 300 g/per test, the dosage content of reducing agent was 8%, roasting time was 60 min. The detailed roasting results at 700 ~ 850°C are shown in Fig. 3, error bar is standard deviation of no less than 3 separate tests. From Figure 3, it was found that roasting temperature had significant influence on both Fe recovery and Fe content of the concentrate, and that Fe content (Figure 3b) gradually increased as the temperature rose from 700 to 850 °C.
To investigate roasting time, the roasting temperature was kept at 800 °C, and other conditions were kept the same as above, with the time varying from 40 ~ 80 min (Figure 4). It was observed that as the roasting time increased, the Fe recovery (Figure 4a) at first increased, and then decreased slightly; however, the Fe content (Figure 4b) in the final concentrate increased gradually for the time range of 50 ~ 80 min.
This is because the mineral particles are controlled by the chemical reaction, which is in turn governed by the diffusion efficiency involved in the roasting process, when the roasting time was short; the reaction process was not completed. On the other hand, if the roasting time was too long, the Fe mineral was deoxidized, leading to the formation of a weakly magnetic FeO or Fe3O4-FO complex , which would decrease Fe recovery as is shown in Figure 4a.
3.3 Quantitative XRD analysis of the reduction roasting process
To investigate the effect of magnetic separation  (which was based on magnetic roasting) for oolitic iron ore, it was necessary to examine the mechanism of the roasting process from microscopic view, especially the transformation of Fe mineral phases and other components during the chemical reaction.
The fixed roasting condition was roasting time 60 min, reducing agent 8%. XRD patterns at 600 ~ 900 °C are shown in Figure 5, it can be easily seen that the content of Fe2O3 decreased with temperature, and that Fe3O4 increased with the rise of temperature (up to 800 °C). The detailed mineral components and content at various temperatures are listed in Table 5, these results are not obtained by merely a common XRD analysis (such as a common XRD pattern). Techniques such as quantitative XRD analysis, identification, & analysis of ore-minerals were also applied. Together with Figure 5, it can be seen that Fe2O3 content decreased, but Fe2O3 existed even at a high temperature of 900 °C, indicating the reduction was still incomplete. The crystal grain size was evaluated using the conventional Scherrer formula, and its value was ~0.3-0.9 μm.
Fe3O4 (magnetite) increased considerably to maximum at 800 °C, with a content of ~64%, however, at 900 °C it decreased to ~41%. From 600 to 700 °C, FeO was not observed in the roasted product, suggesting that there was no over-reduction. When the roasting temperature increased to 800 °C, there was a small amount (4.3%, Table 5) of FeO generated, which means over-reduction happened in a slight degree.
At 900 °C, content as high as 23.6% FeO (in the form of Wustite) was generated, indicating large over-reduction, this accounts for the decrease of Fe3O4 at 900 °C, which is also in agreement with classical thermodynamics  (below, Eq. 1-3). Due to the overly increased FeO, Fe3O4 melted into FeO, forming a type of weakly magnetic solid-melt-body; this can worsen the effectiveness of weak magnetic separation.
When the temperature was less than 800 °C, the roasted product contained nearly no Fe2SiO4 (Ferric silicate, weakly-magnetic). However, after the roasting temperature increased to 900 °C, the content of Fe2SiO4 increased significantly to 8%. This hugely degraded the final roasting product’s quality, as Fe2SiO4 could decrease the separation efficiency.
With the increase of roasting temperature, the content of Ankerite decreased, CaMg0.32Fe0.68(CO3)2 in the form of Ankerite was decomposed to other minerals at 700~800 °C, e.g., CaCO3, this is the reason for the increase of Calcite (CaCO3). However, when the temperature reached 900 °C, the content of Calcite decreased considerably, indicating CaCO3 was decomposed at that high temperature.
The content of Ca5(PO4)3F (Fluorapatite) increased with the increase of temperature, this suggests that phosphorous was mainly present as the high melting point Ca5(PO4)3F phase. Si was mostly present as high melting point Quartz, and weakly magnetic Ferric silicate (Fe2SiO4). The roasted product Fe3O4 cannot react with SiO2, however, with the introduction of CO, reaction could happen and 2FeO·SiO2 can be generated, which would decrease the efficiency of magnetic roasting.
According to Fe-O thermodynamics data and equilibrium diagram [25, 26], α-Fe2O3 can react with reductive gas phase CO in the roasting process, and reaction would follow the sequential steps: 3Fe2O3 → 2Fe3O4 (Eq. 1) → 6FeO (Eq. 2) (shown in below equations).
At ~843 K, the Gibbs free energy ΔrGm between the reaction of SiO2 and the generated Fe3O4 is positive, and will increase with increasing temperature, indicating that Fe3O4 would not react with SiO2 and can exist steadily at this temperature range.
However, with the introduction of CO, ΔrGm among Fe3O4, SiO2 (Eq. 3) and CO can be negative, this may cause the generation of 2FeO·SiO2 and loss of Fe3O4
Table 6 lists each iron mineral phase and its distribution, and it was found that the magnetite phase increased remarkably from 600 to 700 °C, and reached the peak value of ~95% at 800 °C. Then, it decreased somewhat to ~88%, however, the total Fe increased, which suggests that part of magnetite was transformed to Wustite, this trend is in consistent with the quantitative analysis of XRD patterns.
The hematite phase decreased steadily as the temperature increased; however, at 900 °C there was a slight increase, XRD analysis (Table 5) showed that its content (mass) increased from 3.5% to 3.6%, this may result from the lack of reducing agent. The total Fe was relatively low, as the ash residue from the roasting of the reducing agent (coal powder) stayed within the final roasted product.
3.4 Particle size distribution on magnetic roasting result
The particle distribution of feeding has significant influence on the magnetic roasting efficiency; generally, the roasting of Fe2O3 was governed by the diffusion process. Particle sizes less than 2.0 mm were investigated; the separation results after magnetic roasting are listed in Table 7.
From Table 7, ignoring the effect of crystal size, the size of 0.5 ~ 1.0 mm seems to be suitable for roasting and it can be revealed that feeding size has important role in the final product’s quality. When the size is big, the efficiency of heat and mass transfer through the core of mineral in the roasting process is poor, while for very small size, the permeability within the reaction system is low.
3.5 Reductive characteristics of oolitic iron ore during phase change
The metallic mineral within the oolitic iron sample is mainly in the form of hematite; due to its varying constituents, the reduction reaction occurred around the outer circle of the oolitic particles more or less at some degree. As in Figure 6, the reduction area was 3/5 (Figure 6a) of the oolitic sample, and full reduction to magnetite (Figure 6f) could be seen.
Please note, that for example in Figure 6a, the original oolitic particle was almost all hematite (Fe2O3), then after roasting, the outside area was changed into magnetite (Fe3O4), the color of hematite is lighter. Also in Figure 6d, the outer circle of the oolitic particle was slightly darker than that of inner core, and an interesting phenomenon is that all magnetite is observed to sit on the outside of the oolitic shape, which suggests that the reduction reaction occurs from out to inside during the roasting process.
Using the reduction area data (ratio 0~5/5) of the roasted samples, the detailed statistical values and distribution are presented in Table 8, for the condition of 800 °C and 60 min roasting. It can be found that,
from the analysis of the microstructure by microscope, the roasting process did not change the oolitic structure, i.e., after roasting the embedding relationship among the minerals is similar to that of the original oolitic iron ore, with the mineral phase changing from Fe2O3 to Fe3O4.
From quantitative analysis of the reduction area in Table 8, it is revealed that the reduction reaction was relatively well-distributed at each particle degree, from inner to outside.
At the same reduction area ratio 3/5, for fine particles with size < 0.3 mm, the dense value is 71.83% (from Table 8), while for the case of size 0.3 ~ 0.5 mm, the value is 47.46%. This suggests that fine particles have a higher degree of reduction, and coarse grade cannot be easily magnetically recovered.
3.6 Reduction-magnetic beneficiation with industrial scale application
3.6.1 Initial magnetic-roasting test using low content reductant
In commercial industry for beneficiation practice [27, 28], sometimes the feeding quality of the final pilot-scale (or at real production stage) is different (usually lower) from the original ore in lab-scale. To mimic real industrial iron beneficiation, so as to optimize and facilitate the process control, 5% (mass) surrounding rock was added to 400 Kg iron ore, and two types of coal (SP coal, and WG coal) were applied. The mixed iron ore sample (denoted as SP) had the chemical compositions (mass, %) as follow: TFe (40.51), CaO (10.49), MgO (1.95), SiO2 (12.02), Al2O3 (4.97), P (0.72), S (≤0.1). The recoverable Fe in this type of SP sample is mainly in the form of hematite-limonite (89.19%).
To preliminarily test the effect of magnetic roasting, the following parameters were applied: roasting temperature 700 ~ 900 °C, roasting time 40 ~ 70 min, dosage content of reducing agent 3% ~ 11% (the content is 8% in the mechanism study Section 3.2). Some results can be seen as in Table 9, others are in Appendix A. The production yield of Middling 2 (intermediate product) is not low enough, this suggests that there are still small amounts of weakly magnetic ore which would affect the quality of the concentrate.
3.6.2 Enhanced magnetic-roasting and magnetic separation industrial test
The initial magnetic roasting was performed with the dosage content of 3% ~ 11% reducing agent, and roasting time 40 ~ 70 min. However, the results suggested that the magnetic reduction process did not react completely, and there is still some Fe which was not turned into Fe3O4.
Hence, the content of reducing agent was increased from 3% ~ 11% to that of 11%, 15% and 20%, and the roasting time was prolonged to 75 min, 90 min and 120 min. Other conditions were kept as above, i.e., temperature 800 °C, magnetic field : 1.5A (1500 Oe ~ 1600 Oe) 0.8A (800 Oe ~ 850 Oe), particle size -2.0 mm, WG coal and SP coal. The magnetic roasting results of SP samples, using WG coal and SP coal, are presented in Table 10, and Table 11, respectively.
From Table 10, when the roasting time was increased from 90 min to 120 min, the roasting effect was greatly improved, see SP-74 (in bold), Fe content was up to 56.9%. The Fe recovery was 83% ~ 85%, and sometimes as high as 87%, the overall magnetic roasting results are quite satisfactory. Also from the comparison of SP-78 and SP-79 in Table 11, the higher content of reducing agent together with increase of roasting time brought in a better product quality for SP-78.
From Table 10 and Table 11, the magnetic roasting results showed no apparent difference from using either WG coal or SP coal, when the heating efficiency or the effect of ash on afterward grinding is not considered. The reduction process of magnetic roasting can be achieved by any type of coal as reductant, however, SP coal would release more high-sulphur pollutants, which will worsen the surrounding environment.
This study provides a systematic investigation of the quantitative analysis of reductive magnetic-roasting processes for oolitic iron ores, from both a microscopic view and industrial-scale application. Firstly, by mineralogical study, the raw iron particles are shown to be mainly in oolitic shape, composed of hematite and clay minerals with a circling layered structure, with siliceous quartz or clay like kaolinite in the middle core. The roasting conditions were studied, Fe content was found to gradually increase as the temperature increased. The most important contribution of this research is the quantitative study of reduction roasting processes and reductive characteristics during phase changes, using XRD analysis, and distribution of reduction area, the following main observations and conclusions can be drawn:
During the roasting process, the content of Fe2O3 decreased with temperature. Fe3O4 (magnetite) increased considerably from 600 to 800 °C, to maximum at 800 °C (~64%), however, at 900 °C it decreased to ~41%.
From 600 to 700 °C, FeO was not observed, indicating no over-reduction. When the temperature increased to 800 °C, a small amount (4.3%) of FeO was generated, which means there is over-reduction. At 900 °C, 23.6% FeO was generated, indicating large over-reduction; this is the reason for the decrease of Fe3O4 at 900 °C, which is also in accordance with classical thermodynamics.
Fe3O4 melted into the overly increased FeO, which formed a type of weakly magnetic solid-melt-body, and this can worsen the efficiency of weakly magnetic separation.
When the temperature was below 800 °C, there was no Fe2SiO4 (Ferric silicate, weakly-magnetic) generated. However, after temperature increased to 900 °C, Fe2SiO4 increased significantly to 8%, this would degrade the final roasting product’s quality as Fe2SiO4 could not be recovered.
All magnetite (Mt) was observed to be at the outside of oolitic particle, indicating that the reductive reaction occurred from outside to inside during the roasting process.
The original oolitic structure and embedding relation among the minerals did not change after roasting, with mineral phase changing from Fe2O3 to Fe3O4. The reduction reaction was relatively well-distributed at each particle distribution, and coarse grade was not easily reduced.
Finally, to optimize real industrial beneficiation, the content of reducing agent was increased to 11%, 15% and 20%, with roasting time to 75 min, 90 min and 120 min. The overall magnetic-roasting results were satisfactory at the current stage, Fe content reached ~56.9%, Fe recovery was 83% ~ 85%, the two types of coal showed no significant difference. This research would help commercial development of such refractory iron ores, future work can be reverse flotation to upgrade the concentrate.
This research was supported by the National Natural Science Foundation of China (grant numbers 2013BAB03B03, 51674207, 51304162), Doctoral Foundation of Southwest University of Science and Technology (17zx7161).
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About the article
Published Online: 2017-12-29
Conflict of interest: Authors state no conflict of interest.
Citation Information: Open Chemistry, Volume 15, Issue 1, Pages 389–399, ISSN (Online) 2391-5420, DOI: https://doi.org/10.1515/chem-2017-0043.
© 2017 Tiefeng Peng et al.. This work is licensed under the Creative Commons Attribution-NonCommercial-NoDerivatives 4.0 License. BY-NC-ND 4.0