Description:FORM 2
THE PATENT ACT 1970
(39 OF 1970)
&
The Patent Rules, 2003
COMPLETE SPECIFICATION
(See Section 10 and Rule 13)

1 TITLE OF THE INVENTION :
A PROCESS FOR UPGRADING THE IRON CONCENTRATION IN A IRON ORE WITH MANGANESE MINERAL BY SULPHUR REDUCTION ROASTING FOLLOWED BY ACID LEACHING.



2 APPLICANT (S)

Name : JSW STEEL LIMITED.

Nationality : An Indian Company incorporated under the Companies Act, 1956.

Address : JSW CENTRE,
BANDRA KURLA COMPLEX,
BANDRA(EAST),
MUMBAI-400051,
MAHARASHTRA,INDIA.




3 PREAMBLE TO THE DESCRIPTION

COMPLETE

The following specification particularly describes the invention and the manner in which it is to be performed.

FIELD OF THE INVENTION

The present invention relates to the development of an extractive metallurgy process for the removal of manganese values from an iron rich ore contaminated with Mn values less than 10%. The main objective of the study is to enrich the Fe content and simultaneously remove Mn impurity. More particularly, the present invention is directed to a process for removal of manganese content in low grade iron ore fines and proportionately increasing the Fe content in the final product and the feed material comprising Fe grade varying from 53 to 55% and Mn content in the ore varying between 5 and 9 %. More particularly, the ore is roasted with sulphur reductant, where a reaction between S and Mn values takes place preferentially due to the affinity of S for Mn. The reduction reaction is enabled in a temperature range between 550 and 650 oC and the reduction time less than 1 hour. The Mn and S rich reacted mass is preferentially leached with a dilute acid treatment for leaching times less than 45 min, which enables removal of Mn from the ore to the leach liquor while the Fe values get enriched to 59 and 62% Fe suitable for iron ore pellet. The residual Mn in the ore mass reduces to less than 1% Mn with a Mn removal efficiency greater than 90%.

BACKGROUND OF THE INVENTION

Low grade iron ores with Mn content between 5 to 9% are less preferred ore for iron extraction, although Mn alone is a valuable resource for steel making. The contamination of Mn in iron ore seriously affects the quality of pig iron in blast furnace. The subsequent steel making process with higher Mn in hot metal has several process inefficiencies. The high Mn content removal requires higher oxygen, refractory and lime consumption. The slag volumes generated are larger and affects the refractory consumption, higher power input and lower productivity. Hence, Mn ores with 5-10% is less preferred for use and is available at low costs.

Bhomman Iron Ore Mine of M/s. JSW Steel Limited., Chitradurga consists of total reserves of 55.9 million tonnes of iron ore. Fe content in the feed material is varying from 53 to 55% and the Mn content varies between 5.00 and 9.0. The hot metal produced in the blast furnace, using such lean Mnrich iron ore leads to several process abnormalities in steel making. Removal of manganese from low grade iron ore fines is essential before using in iron making and steel making process.

Beneficiation of the iron ore to remove preferentially Mn based minerals in iron ore by conventional beneficiation technique is one manner where Fe values are being upgraded. The processes adopted in conventional beneficiation depends on the physical and chemical characteristics such as specific gravity, magnetism, surface characteristics etc which enables separation of Mn values from Fe values. Such beneficiation processes use several unit processes such as washing, gravity concentration magnetic separation, floatation, heavy media separation etc. As the Mn minerals in the iron ore has similar physiochemical characteristics of iron ore, the conventional beneficiation does not result in significant Fe upgradation and recovery.

PRIOR ARTS

In the prior art there are several techniques to selectively remove Mn values from lean iron ore with Mn values.

Abou-El-Sherbini, studied the selective Mn removal from a low grade pyrolussite Mn ore where the Mn values were recovered by reaction of the ore with elemental sulphur at a temperature between 300 and 400 oC followed by leaching with water of dilute sulphuric acid. The Mn rich liquor is concentrated and Mn was precipitated as MnO2 with ammonium peroxidisulphate or electrolysis on Pt anode. The original ore was 27.13% Mn in it. While the present ore in the present innovation had 6 to 8 %Mn in the predominantly iron ore, the ore in prior art is rich in Mn content. The primary objective of the present investigation is to remove Mn values so that iron values are significantly enriched. The present ore characteristics and reactivity are different and the reduction conditions are unique to the present ore. The iron recovery is more valuable than the Mn recovery in the present study.
Kh.S. Abou-El-Sherbini, Separation and Purification Technology 27 (2002) 67–75

Yongjie Liu et al. have reported a novel Process of Reduction Roasting Manganese Ore withSulfur Waste and Extraction of Mn by Acid Leaching. The process involved reduction of Mn rich ore with 34.75% MnO2 and 41% Hematite along with 8.32% silica and 5.15 % Alumina as major gangue material being roasted with Sulphur to convert MnO2 to MnS. The reduction conditions were optimized at 450 oC for 30 min holding time;theSulphur to MnO2 molar ratio was maintained at 0.40;leaching by deionized water gave 24.3 wt%, while it is above 95% when leached with a 1 molar sulphuric acid. Where as the primary objective of the prior art is to extract Mn values, the present invention relates to Sulphur removal from an ore where MnO2 to 6 to 8% , which was removed to upgrade the Fe values in the ore to 62%
Yongjie Liu, Fupeng He, Donglai Ma, Qingqing Hu and Zhixiong You, Metals 2022, 12, 384. https://doi.org/10.3390/met12030384

Yuanbo Zhang et al., have used a Mn rich ore with 21% Mn along with 10%Fe and 6.9 % Al2O3 , 31.62% SiO2 and 2.25%CaO. to win the Mn values using sulpur reduction roasting followed by acid leaching to recover Mn values in the ore. The reduction roasting temperature was optimized at 550 oC with a S/Mn ratio of 0.5 and leaching with 1 molar sulphuric acid for 5 min within solid to solution ratio of 1:5. Compared to the prior art where Mnextraction was targeted, the present invention relates to upgrading an iron ore with lean Mn content, where Mn was 6 to 9% and it is selectively removed to enrich the Iron oxide that can be used for Iron making primarily while Mn values may be concentrated and concentrated to MnO2. The Mn values are too low compared to the prior art and the gangue association are also different. Kaolinite and Gibbsite is seen in the invented process while it was quartz in the prior art.
Yuanbo Zhang, Zhixiong You, Guanghui Li, Tao Jiang, Hydrometallurgy 133 (2013) 126–132

Lin Deng et al., studied a 30%Fe- 4%Mn containing pyrolusite adsorption residue for the recovery of Fe and Mn values by reacting the residue with ammonium sulphate at 200 to 350 oC followed by 0.5 M sulphuric acid leaching. While this prior art uses ammonium sulphate the invented process uses pure sulphur for reduction.
Lin Deng, Bing Qu, Shi-Jun Su, Sang-Lan Ding and Wei-Yi Sun, Metals 2018, 8, 38; doi:10.3390/met8010038

Zhixiong You et al., have suggested the extraction of Mn from Mn rich ore by selective sulphate roasting by treating the ore with (sulphur dioxide + N2) gas mixture at 500 oC for 60 min at a sulphur dioxide partial pressure between 0.5 to 1.0 followed by water leaching at 50 oC for 15 min with solid to aqueous solution ratio of 1:5 to selectively remove Mn as a sulphate in aqueous solution. Whereas the prior art talks about a gas based reduction, the invented process is about direct solid-solid reaction between S and the ore.
Zhixiong Youet.al., Hydrometallurgy, 156, 2015, 225-231,
Z. You, G. Li, , Z. Peng, L. Qin, Y. Zhang, T. Jiang, J. Min. Metall. Sect. B-Metall. 53 (2) B (2017) 115 - 122

The Patent No. CN20121268561 discloses a reduction process involving tunnel kiln to reduce a high Mn ore with different sulphur pre heated at 310 oC for 6 min followed by reduction in air free condition at 700 oC for 10 min. A 22% sulphuric acid could leach 94.4% Mn in the ore. While the prior art deals with ore rich in pyrolusite, psilomelane, hausmannite and brunite, the present invented process deals with a Fe rich ore where Mn based mineral is low and the gangue contents are low.
JIANG TAO ZHANG, YUANBO LI, GUANGHUI FAN, XIAOHUI YOU, ZHIXIONG GUO,YUFENG HUANG, ZHUCHENG YANG, YONGBIN LI, QIAN CHEN, XULING ZHOU, YOULIAN DUAN, DAOXIAN LI, ZHENZHEN LUO, WEI SU ZIJIAN, CENTRAL SOUTH UNIV

The patent CN201410024804 discloses, a wet reducing leaching method for manganese oxide mineral, is characterized in that, by sulfur based reducing agent and manganese oxide mineral with mass ratio 0.05 ~ 0.5:1 in temperature is first to stir in the neutral aqueous media of 60 ~ 95 DEG C reduction reaction occurs; After reduction reaction completes, filtering separation, gained filter residue concentration is that the sulphuric acid soln of 0.5 ~ 2.0mol/L leaches at 40 ~ 90 DEG C, and solid-liquid separation, obtains manganese leach liquor; Described sulfur based reducing agent is one or more in ammonium sulfide, sulphur hydrogenation ammonia, sulfurated lime, calcium sulfhydrate, sodium sulphite, Sodium sulfhydrate.The primary purpose of the prior art patent is to recover Mn from the Mn rich ore. The invented process is dealing with a predominantly iron ore with a minor fraction of Mn which is removed by sulphur based reductant. The minerals starting and the final objective of the invented process targets iron in the ore. The gangue associated are different.
ZHONG HONG LI, CHANGXIN WANG, SHUAI CAO, ZHANFANG
CENTRAL SOUTH UNIV

Yunhyua Wang et al. report a high temperature reaction of manganese oxide ore, ferrous sulfate and water , in the ratio 1:3:20 or with manganese oxide ore, sulphur and water the reaction is carried out at 120 to 350 oC for 30 min followed by washing. To separate Mn as MnSO4 . While this prior art process involves reduction reaction with water as an ingredient, the invented process uses only reaction between the ore and Sulfur. The target in the prior art is to recover Mn from the ore, the invented process Mn is a by product, which has to be further concentrated for effective recovery while the major focus is to get iron ore concentrate enriched to 58 to 62% Fe content. YUNHUA WANG, YONGCUN LAI, Patent No. CN100567167C

From the above disclosed state of prior art, it is understood that the present invented process is intended to upgrade the iron ore with Mn contamination rather than other studies which focus on Mn values in Mn rich ores. The association of Mn values with gangue is different from that in high Mn ore.

The present invention targets the development of a process where Mn is effectively removed from the ore by a simple extractive metallurgy process technique. The ore so produced with about 1% residual Mn can be effectively blend and pelletized for use in iron blast furnace. In this invention, a sulphur based reduction roasting is carried out to preferentially react Mn with S at appropriate reaction condition. This is followed by a chemical leaching of the reacted mass where the Mn product of reaction with S is leached with dilute acid to transfer the Mn values from ore to aqueous phase. As the Mn is leached, the Iron ore gets upgraded to >60% Fe. The Mn values that is leached is concentrated and can be concentrated as MnO or Mn metal in subsequent processes. The upgraded iron ore fines could be used for pellet making for use in iron blast furnace and Mn in aqueous solution for Mn recovery. Theoretically, for every 1 ton of iron ore processed 60 kg of Mn values (77 kg MnO) is produced and Fe recovery will be varying between 80 to 84%

OBJECTIVES OF THE INVENTION

1. The basic object of the present invention is directed to a process for upgrading Fe content in iron ore by removal of Manganese values from iron rich ore contaminated with Mn values less than 10 %.
2. A further object of the present invention is directed to said process to ensure that the iron ore with 53 to 55% Fe and 5 to 9% Mnvalue initially is suitably reduction roasted with sulphur followed by dilute acid leaching to obtain two products.
3. A still further object of the present invention is directed to said process wherein the target product is the iron ore concentrate which is enriched from 55 to 62% Fe and the iron recovery is greater than 80% and the residual Mn content <1% and wherein the second product is the Mn rich aqueous leach liquor where Mn values are concentrated to extract MnOwhich can be used for Mn extraction. Experimetally, for every one ton of initial ore >800 kg of Iron rich concentrate is produced and 71 kg of MnO is extracted.

4. A still further object of the present invention is directed to development of suitable reduction roasting condition to ensure that the ore is reacted with S for recovery of Mn, involving appropriate temperature and time optimization for sulphur roasting process.

5. Another object of the present invention is directed to said process wherein the reacted mass is leached with required concentration of dilute mineral acid that removes the Mnvalue and at the same time it enriches the Fe values of the ore suitable for iron ore pellet making.

6. Yet another object of the present invention is directed to said process so designed as to ensure > 90% of original Mn is removed from the ore to give an iron concentrate of 58 to 62% Grade and greater than 80% Fe recovery in concentrate and the final concentrate has less than 1% Mn.

7. A further object is directed to said process providing means to concentrate the leach liquor for pure MnO recovery which can be a product to be sold for ferroalloy industry.

SUMMARY OF THE INVENTION

The basic aspect of the present invention is directed to a process for upgrading Fe content in iron ore by removal of Manganese values from iron rich ore contaminated with Mn values less than 10 % comprising :
subjecting the said iron ore having Mn values to reduction roasting with sulphur as a reductant at temperatures of 550 to 600oC for a period of 60 minutes ; followed by
leaching the thus reduced product with dilute sulphuric acid whereby the iron ore having initial iron content 53-55 % to 58 -65% Fe and reducing Mn content from said less than 10% to upto about 1%.

A further aspect of the present invention is directed to said process wherein said iron rich ore comprises
predominantly iron rich ore fines 53 to 55 % hematite; 16 to 18 %goethite; 3 to 4 % magnetite and 9.5 to 10.5% pyrolusite, along with gangue minerals kaolinite 2.5 to 3.5%, quartz 2.7 to 3.6% and others 2.5 to 3.0% and the said ore have a particle size range of 80 to 85% -0.150µm, 78 to 80% -75µm, 72 to 76% -45µm, 60 to 64% -10µm and 25 to 28% -2µm.

A still further aspect of the present invention is directed to said process wherein said iron ore having Mn values is intimately mixed with sulphur as a reductant, having a minimum purity of 90% preferably in the range of 90% to 95 % and the reductant sulphur in the range of less than 15% preferably in the range of 10 to 15% and the entire mass is subsequently briquetted.

A still further aspect of the present invention is directed to said process wherein said iron ore raw material and the reductant including said briquetting were heated at temperature ranging between 500 and 600 oC in a suitable reactor for a period of about 60 min to enable reaction between Mn values in the ore and the sulphur reductant and cooled to room temperature and powdered to 28oC preferably about -0.5 mm powder.

Another aspect of the present invention is directed to said process wherein said iron ore raw material and the reductant including said briquetting carried out under reduction condition and the reduced product is subjected to an acid leaching treatment in a dilute about 2 Molar sulphuric acid at a solid to acid ratio maintained constant at about 100 ml/gm of reduced mass, and wherein the mass is agitated in a stirred reactor at about 200 rpm for leaching duration varying between 15 and 25 min., said leach liquor being filtered and washed in water thrice to separate iron oxide concentrate solids from leach acid liquor.

A further aspect of the present invention is directed to said process wherein said leaching process is carried out such as to achieve a final iron concentrate with Fe grade between 58.5 to 61.6% and the Mn content <1% and Fe recovery above 82 to 90%.

A still further aspect of the present invention is directed to said process carried out with said finally iron concentrate and Mn values which enabled 58.5 to 61.6% Fe concentrate grade with an Fe recovery of 82 to 90% and with a Mn removal efficiency above 90%

The above and other aspects and advantages of the present invention are described hereunder with reference to the following accompanying non limiting illustrative drawings and example.

BRIEF DESCRIPTION OF THE ACCOMPANYING DRAWINGS

Figure 1: Micrographs of iron ore fines.
Figure 2: Process flow diagram to upgrade Fe ore and recover Mn values.
Figure 3(a)-(c): show the effect of excess sulphur over stoichiometry at three different leaching time on the Fe grade, Fe recovery and concentrate recovery.
Figure 4 (a)-(c): show effect of excess sulphur over stoichiometry at three different leaching time on the Mn loss in ore, Mn in leach liquor, and Mn% in concentrate.

DETAILED DESCRPTION OF THE INVENTION WITH REFERENCE TO THE ACCOMPANYING DRAWINGS AND EXAMPLES

The present invention relates to the development of an extractive metallurgy process for the removal of manganese values from an iron rich ore contaminated with Mn values less than 10%. The main objective of the study is to enrich the Fe content and simultaneously remove Mn impurity.

The as-received iron ore fines from Bhomman Mines of JSW Steel Limited was subjected to chemical analysis using XRF (X-ray fluorescence). Chemical analysis of the iron ore is shown in Table 1.

The received material was ground to 80 to 85% passing 150 µm (100#) for reduction roasting – leaching studies to recover both iron bearing minerals and manganese. The size analysis of the ground material is shown in Table 2. From size analysis it was found that -150 µm varying from 80 to 85%, -45 µm varying from 72 to 76% and -10 µm varying from 60 to 64%. The as received iron ore sample was subjected to X-ray diffraction (XRD) analysis to know the different mineralogical phases present in the sample. Table 3 shows the XRD phase analysis. The XRD analysis indicated that iron bearing minerals are goethite, hematite and magnetite. Goethite is the major iron bearing mineral. The manganese bearing mineral is pyrolusite. The gangue bearing minerals are kaolinite and quartz. The micrographs of iron ore samples are shown in Figure 1.

Table 1: Chemical analysis of iron oreby XRF
Element Wt.%
Fe 53.0 to 55.0
Mn 5.00 to 9.00
SiO2 2.80 to 3.20
Al2O3 1.50 to 2.00
S 0.01 to 0.04

Table 2: Size analysis of iron ore fines
Size, µm %
-150 80 to 85
-75 78 to 80
-45 72 to 76
-10 60 to 64
-2 25 to 28

Table 3: Phase analysis of the feed iron ore with manganese by XRD

Phase Chemical Formula Percentage (%)
Hematite Fe2O3 53.0 to 55.0
Goethite FeO(OH) 16.0 to 18.0
Kaolinite Al2Si2O5(OH)4 2.5 to 3.5
Pyrolusite MnO2 9.5 to 10.5
Magnetite Fe3O4 3.0 to 4.0
Quartz low SiO2 2.7 to 3.6
Others 2.5 to 3.0

The aim is to develop simple cost effective process to remove manganese from low grade iron ore fines and proportionately increasing the Fe content in final product. The process followed for removing the manganese from low grade iron ore fines is shown in Figure 2.

Lab Scale Experiments:

Initial studies were carried out in 300 gm batch size of iron ore. The iron ore was mixed with different ratio of S content. The sulphur reductant was maintained at stoichiometry, 5 and 10% excess the stoichiometric requirements. The ore mixed with S powder was briquetted. The briquette was placed in a iron crucible. The briquette was roasted in a muffle furnace at three different temperatures 500, 550 and 600oC. At each temperature, the roasting time was maintained at 60 min. The furnace was well vented to remove SO2, if any, formed during reaction. The roasted sample was cooled in air to room temperature. The briquettes were crushed and screened using -0.5mm powder.

Leaching studies were carried out in 50 g potion of reacted powder mass. The powder was taken in a beaker with dilute sulphuric acid of concentration 2 mol/L solution concentration and maintaining in 500 ml solution. The powder with the acid was agitated using a stainless steel stirrer whose rpm was kept at 200. The leaching time was varied at three different levels 10, 15 and 20 min. After completion the leaching process for the set time, the solution was filtered. Then, the leached residue was washed three times using distilled water and dried for chemical analysis. The residue obtained after leaching is considered as final product. The weight of the residue was taken for weight recovery calculation. The leach liquor was concentrated and the Mn values were examined. Other options are available and those are aqueous sulfur dioxide, hydrogen peroxide, organic biomass reductants, and biological bacteria. Usage of these reductants leads to low leaching effeiciency and high production cost, hence not considered.

The experimental details are shown in Table 4 with regard to the reduction roasting trials. The test results of leaching experiment are shown in Table 5. The reduction in MnO at optimum test conditions where the concentrate has less than 1% is shown in Table 6. The Figure 3(a)-(c) gives the stoichiometric excess of S as a function of the Fe grade and Fe values in the concentrate as a function of leaching duration. The Figure 4(a)-(c) shows the loss in Mn from the ore due to reaction with S and the Mn values in the leach liquor, which can be used for recovering the Mn value as a function of excess S taken over stoichiometric values and leaching duration.

Table 4: Experimental conditions used in the study

Variables Parameters
Sulfur,% 5, 10, 15
Roasting Temperature, oC 500, 550 and 600
Roasting time, min 60
Leaching time, min 15, 20, 25
Leaching temp.oC 28
Mixing (stirrer), rpm 200

Table 5: Test results at different parameters

Table 6: Test results at optimum parameters

Description Feed Product
Sulfur,% 10 to 15
Roasting Temperature, oC 550 to 600
Leaching time, min 20 to 25
Iron Recovery,% 82 to 90
Fe,% 53.0 to 55.0 58.5 to 61.6
Mn,% 4.00 to 7.00 < 1%
SiO2,% 2.80 to 3.20 3.68 to 4.56
Al2O3 1.50 to 2.00 1.38 to 1.96
S,% 0.01 to 0.04 0.02 to 0.05
The manganese oxides are changed from high valence to low valance compounds during reduction roasting process as follows:
MnO2 ---> Mn2O3 ---> Mn3O4 ----->MnS (or MnSO4)
The Major chemical reaction evolved during the leaching process for removal of manganese and extraction of iron from the roasted product is as follows:
MnS + H2SO4 ----> MnSO4 + H2S

Thus, through the reduction roasting followed by acid leaching process,the primary product Fe 58.5 to 61.6% with weight recovery 82 to 90%. The MnO was reduced from 7 - 9% to <1.0%. making the material directly useful for blast furnace pellet feed.

The % Mn in iron rich residue <1.0% with higher Fe content is considered as optimum results and the test conditions with respect to these results are considered as optimum test conditions.

, Claims:We Claim:
1. A process for upgrading Fe content in iron ore byremoval of Manganese values from iron rich ore contaminated with Mn values less than 10 % comprising :
subjecting the said iron ore having Mn values to reduction roasting with sulphur as a reductant at temperatures of 550 to 600oC for a period of 60 minutes ; followed by
leaching the thus reduced product with dilute sulphuric acid whereby the iron ore having initial iron content 53-55 % to 58 -65% Fe and reducing Mn content from said less than 10% to upto about 1%.

2. The process as claimed in claim 1 wherein said iron rich ore comprises
predominantly iron rich ore fines 53 to 55 % hematite; 16 to 18 %goethite; 3 to 4 % magnetite and 9.5 to 10.5% pyrolusite, along with gangue minerals kaolinite 2.5 to 3.5% , quartz 2.7 to 3.6% and others 2.5 to 3.0% and the said ore have a particle size range of 80 to 85% -0.150µm, 78 to 80% -75µm, 72 to 76% -45µm, 60 to 64% -10µm and 25 to 28% -2µm.

3. The process as claimed in anyone of claims 1 or 2 wherein said iron ore having Mn values is intimately mixed with sulphur as a reductant, having a minimum purity of 90% preferably in the range of 90% to 95 % and the reductant sulphur in the range of less than 15% and the entire mass is subsequently briquetted.

4.The process as claimed in anyone of claims 1 to 3 wherein said iron ore raw material and the reductant including said briquetting were heated at temperature ranging between 500 and 600 oC in a suitable reactor for a period of about 60 min to enable reaction between Mn values in the ore and the sulphur reductant and cooled to room temperature and powdered to-0.5 to -0.3 preferably about -0.5 mm powder.

5. The process as claimed in anyone of claims 1 to 4 wherein said iron ore raw material and the reductant including said briquetting carried out under reduction condition and the reduced product is subjected to an acid leaching treatment in a dilute about 2 Molar sulphuric acid at a solid to acid ratio maintained constant at about 100 ml/gm of reduced mass, and wherein the mass is agitated in a stirred reactor at about 200 rpm for leaching duration varying between 15 and 25 min., said leach liquor being filtered and washed in water thrice to separate iron oxide concentrate solids from leach acid liquor.

6. The process as claimed in anyone of claims 1 to 5 wherein said leaching process is carried out such as to achieve a final iron concentrate with Fe grade between 58.5 to 61.6% and the Mn content <1% and Fe recovery above 82 to 90%.

7. The process as claimed in anyone of claims 1 to 6 carried out with said finally iron concentrate and Mn values which enabled 58.5 to 61.6% Fe concentrate grade with an Fe recovery of 82 to 90% and with a Mn removal efficiency above 90%

Dated this the 23rd day of September, 2022
Anjan Sen
Of Anjan Sen & Associates
(Applicants’ Agent)
IN/PA-199