Method for recycling fine iron powder

Abstract

The invention discloses a method for recovering fine iron powder, and relates to the technical field of resource recovery. The invention provides a process route of hydrochloric acid leaching, iron powder reduction, sulfuration impurity removal, evaporative crystallization, size washing and roasting, and impurities such as Cu and Co can be removed through iron powder reduction; then the characteristic that Ksp of NiS and CoS is smaller than that of FeS is utilized, FeS is used for sulfuration impurity removal, and precipitation and recovery of Ni and Co can be completed under the condition that Fe is not lost; finally, through evaporative crystallization, residual trace impurity elements in the impurity removal liquid can be enriched in the solution and are prevented from entering ferrous chloride crystals; and ferrous chloride can be efficiently converted into iron oxide red (or iron oxide yellow) through a roasting process, so that the equipment investment is reduced. According to the method, through optimization of the whole process, the fine iron powder generated by high-grade matte nickel hydrometallurgy can be converted into high-purity iron oxide red (or iron oxide yellow) in a recycling mode, and impurities such as Ni, Cu and Co in the fine iron powder can be effectively recycled.

Description

Technical Field

[0001] The invention relates to the technical field of resource recovery, and in particular to a method for recovering iron ore concentrate. Background Art

[0002] Iron red refers specifically to iron oxide crystals with a specific morphology and particle size. As a transition metal oxide with high chemical and physical stability, alkali resistance, and weak acid resistance, it has been widely used in coatings, building materials, plastics, etc. The chemical formula of iron red is α-Fe2O3, and its color is mostly bright red or dark red. Its hue can be controlled by changing the temperature parameters of the preparation process, and there is a color change rule of yellow-brown-red-purple-black corresponding to low temperature to high temperature.

[0003] At present, the commonly used methods for preparing iron red can be divided into three categories: wet method, thermal decomposition method (also defined as dry method) and wet precipitation and pyrolysis method. Among them, the iron red product prepared by the wet method does not involve high-temperature treatment, so its stability and degree of crystallization are significantly lower than those of the dry method and the dry-wet combined method, but the iron red prepared by this method is brighter and easier to control. Therefore, this method is the most widely used in the industrial production of iron red. The thermal decomposition method can be regarded as the roasting step of the wet precipitation and pyrolysis method, so the iron red prepared by these two types of technology has higher stability, but the brightness of iron red is lower, and there is a tendency to turn into iron black. In the process of industrial production, the raw materials for preparing iron red are usually divided into the following categories: iron-containing waste slag, iron-containing waste liquid, steel mill scrap iron or scrap iron, iron ore, and iron salt. Among them, iron-containing waste slag, iron-containing waste liquid and steel mill scrap iron are the most economical raw materials. Except for the production of special iron red, these raw materials have the best economy.

[0004] During the wet smelting process of high-grade nickel matte, the oxygen pressure iron removal process will produce a large amount of iron concentrate containing a certain amount of valuable elements such as Ni and Cu. The iron concentrate has a low selling price, which leads to the waste of Ni / Cu resources and economic losses. It is also of no benefit to the downstream steel smelting industry. Even Ni / Cu as impurities may pose risks to downstream enterprises. At present, there is a lack of relevant technology for the resource treatment of the iron concentrate, and the efficiency of sulfuric acid leaching of iron concentrate is extremely low, and it is difficult to remove impurities.

[0005] Therefore, there is an urgent need to develop a method that can efficiently and low-cost convert iron ore concentrate into high-quality red iron.

[0006] In view of this, the present invention is proposed. Summary of the invention

[0007] The object of the present invention is to provide a method for recovering iron ore concentrate, aiming to convert the iron ore concentrate into high-quality iron red (or iron yellow) with high efficiency and low cost.

[0008] The present invention is achieved in that:

[0009] In a first aspect, the present invention provides a method for recovering iron ore concentrate, comprising sequentially performing hydrochloric acid leaching, iron powder reduction, sulfidation removal, evaporation crystallization, pulp washing and roasting;

[0010] Hydrochloric acid leaching: The iron ore concentrate is mixed with a leaching solution containing hydrochloric acid for reaction, and then the solid and liquid are separated to obtain a hydrochloric acid leaching solution;

[0011] Iron powder reduction: The hydrochloric acid leaching solution and iron powder are mixed and reacted, and then the solid-liquid separation is performed to obtain the reduced leaching solution;

[0012] Sulfidation removal: The reduced leaching solution is mixed with sulfide for reaction, and then the solid-liquid separation is performed to obtain the reduced liquid;

[0013] Evaporation crystallization: The reduced liquid is heated and evaporated, then cooled and crystallized, and the solid-liquid separation is performed to obtain crude ferrous chloride;

[0014] Washing: Wash the crude ferrous chloride with ferrous chloride solution, and separate the solid and liquid to obtain ferrous chloride crystals;

[0015] Calcination: Calcinate the ferrous chloride crystals, and wash the powder obtained after calcination with water.

[0016] In an optional embodiment, the leaching solution used in the hydrochloric acid leaching process is obtained by mixing hydrochloric acid and ferrous chloride base solution;

[0017] Optionally, the preparation process of the leaching solution includes: mixing concentrated hydrochloric acid with a mass fraction of 30%-40% and a ferrous chloride base solution with a concentration of 2.8mol / L-3.2mol / L, wherein the volume ratio of the concentrated hydrochloric acid to the ferrous chloride base solution is 1:(2.5-3.0);

[0018] Optionally, the mass ratio of iron ore concentrate to leaching solution is 1:(8-12).

[0019] In an optional embodiment, the leaching temperature of the hydrochloric acid leaching is 50°C-70°C, and the leaching time is 1h-4h.

[0020] In an optional embodiment, during the reduction of iron powder, the ratio of the amount of iron powder added to the theoretical amount of iron powder is controlled to be (1.0-1.1):1;

[0021] Optionally, the reaction temperature of the iron powder reduction is 50° C.-70° C., the reaction time is 1 h-3 h, and the pH value of the solution at the reaction endpoint is controlled to be greater than 2.0.

[0022] In an optional embodiment, the process of sulfidation and impurity removal includes: mixing a theoretical amount of sulfide and 5%-10% of a reduction leaching solution, separating the solid and liquid to obtain a slag material, mixing the slag material with the remaining reduction leaching solution, and then separating the solid and liquid to obtain a reduced liquid; wherein the theoretical amount refers to the theoretical amount of sulfide consumed by the nickel and cobalt elements in the total amount of the reduction leaching solution;

[0023] Optionally, the sulfide is selected from at least one of sodium sulfide, ferrous sulfide, and manganese sulfide;

[0024] Optionally, the reaction time with 5%-10% reducing leaching solution is 30min-60min;

[0025] Optionally, the reaction temperature with the remaining reduction leaching solution is 60°C-85°C, and the reaction time is 1h-2h;

[0026] Optionally, the slag is washed and slurried before being mixed with the remaining reducing leaching solution for reaction.

[0027] In an optional embodiment, during the evaporation crystallization process, the evaporation ratio is controlled to be 15%-25%;

[0028] Optionally, the evaporation method is reduced pressure distillation, and the operating pressure is controlled to be 0.005MPa-0.020MPa.

[0029] In an optional embodiment, the temperature is lowered to 30°C-40°C after evaporation, and cooled for crystallization for 3h-5h;

[0030] Optionally, after cooling and crystallization, the solid and liquid are separated, and the obtained filtrate is used as the bottom liquid of the next cycle.

[0031] In an optional embodiment, the crude ferrous chloride is pulp-washed with a ferrous chloride solution having a concentration of 1.5-2.5 mol / L;

[0032] Optionally, during the pulp washing process, the volume of the ferrous chloride solution corresponding to 1 kg of crude ferrous chloride is controlled to be 0.5 L-1.0 L;

[0033] Optionally, the pulping time is controlled to be 10 min-30 min.

[0034] In an optional embodiment, the pulping process includes: mixing crude ferrous chloride and ferrous chloride solution for pulping, then separating the solid and liquid, using the obtained 1%-5% filtrate as the ferrous chloride bottom liquid required for the hydrochloric acid leaching process, and replenishing the remaining filtrate with water as the ferrous chloride solution for the next cycle of pulping.

[0035] In an optional embodiment, the calcination is carried out in a water-containing air atmosphere, the calcination temperature is controlled to be 300° C.-500° C., and the calcination time is 1 h-5 h;

[0036] Optionally, the ferrous chloride crystals are dried and then calcined;

[0037] Optionally, the waste gas generated during the roasting process is introduced into a spray device to obtain a hydrochloric acid solution with a concentration of 1 mol / L to 3 mol / L;

[0038] Optionally, the roasting device used is an air-blast tube furnace, and the air blowing rate is controlled to be 2L / min-3L / min;

[0039] The method further comprises: washing the powder obtained after calcination with water and then drying it.

[0040] The present invention has the following beneficial effects: the present invention provides a process route of hydrochloric acid leaching-iron powder reduction-sulfurization impurity removal-evaporation crystallization-slurry washing-roasting, and impurities such as Cu and Co can be removed by iron powder reduction; then, by utilizing the characteristic that the Ksp of NiS and CoS is less than that of FeS, sulfidation impurity removal can be performed to complete the precipitation and recovery of Ni and Co without losing Fe; finally, by evaporation crystallization, the trace impurity elements remaining in the impurity removal liquid can be enriched in the solution to avoid entering the ferrous chloride crystal; iron red (or iron yellow) is prepared by the roasting process, and ferrous chloride can be efficiently converted into iron red (or iron yellow) without using special equipment (such as spray equipment), thereby reducing equipment investment. The present invention can convert the iron concentrate produced by high-grade nickel matte hydrometallurgy into high-purity iron red (or iron yellow) through resource optimization of the overall process, and can effectively recover impurities such as Ni, Cu, and Co in the iron concentrate. BRIEF DESCRIPTION OF THE DRAWINGS

[0041] In order to more clearly illustrate the technical solutions of the embodiments of the present invention, the drawings required for use in the embodiments are briefly introduced below. It should be understood that the following drawings only show certain embodiments of the present invention and therefore should not be regarded as limiting the scope. For ordinary technicians in this field, other related drawings can be obtained based on these drawings without creative work.

[0042] Figure 1 A process flow chart for preparing red iron using iron concentrate provided in an embodiment of the present invention;

[0043] Figure 2 The following are actual pictures of the products prepared in Examples 1-2 and Comparative Example 1. DETAILED DESCRIPTION

[0044] In order to make the purpose, technical scheme and advantages of the embodiments of the present invention clearer, the technical scheme in the embodiments of the present invention will be described clearly and completely below. If the specific conditions are not specified in the embodiments, they are carried out according to conventional conditions or conditions recommended by the manufacturer. If the manufacturer of the reagents or instruments used is not specified, they are all conventional products that can be purchased commercially.

[0045] The present invention provides a method for recovering iron ore concentrate. Figure 1 , including the following steps:

[0046] S1. Hydrochloric acid leaching

[0047] The iron ore concentrate is mixed with a leaching solution containing hydrochloric acid for reaction, and then the solid and liquid are separated to obtain a hydrochloric acid leaching solution. The metal elements in the iron ore concentrate are leached into the hydrochloric acid leaching solution through hydrochloric acid leaching.

[0048] Specifically, iron ore concentrate refers to a slag with an iron content of more than 55% produced during the production process of the production line, which may come from the hydrometallurgical smelting process of high-grade nickel matte, but is not limited to this.

[0049] In some embodiments, the leaching solution used in the hydrochloric acid leaching process is obtained by mixing hydrochloric acid and ferrous chloride base solution, and the ferrous chloride base solution can come from the subsequent step S5. In the actual operation process, the preparation process of the leaching solution includes: mixing concentrated hydrochloric acid with a mass fraction of 30%-40% and ferrous chloride base solution with a concentration of 2.8mol / L-3.2mol / L, and the volume ratio of concentrated hydrochloric acid to ferrous chloride base solution is 1: (2.5-3.0). By adjusting the amount of concentrated hydrochloric acid and ferrous chloride base solution, the leaching solution can more fully leach the metal elements in the iron concentrate.

[0050] Specifically, concentrated hydrochloric acid can be a commercial concentrated hydrochloric acid product with a concentration of 36%, but is not limited thereto. The concentration of the ferrous chloride bottom solution is close to saturation, and can be 2.8 mol / L, 3.0 mol / L, 3.2 mol / L, etc. The volume ratio of concentrated hydrochloric acid to the ferrous chloride bottom solution can be 1:2.5, 1:2.6, 1:2.7, 1:2.8, 1:2.9, 1:3.0, etc.

[0051] Furthermore, the mass ratio of iron ore concentrate to leaching solution is 1:(8-12), such as 1:8, 1:9, 1:10, 1:11, 1:12, etc. The amount of leaching solution used within this range can fully leach the iron and impurity metals in the iron ore concentrate.

[0052] In some embodiments, the leaching temperature of hydrochloric acid leaching is 50°C-70°C, and the leaching time is 1h-4h. Under this leaching temperature and time conditions, the metal elements in the iron ore concentrate can quickly enter the leaching solution. Specifically, the leaching process can be carried out under stirring conditions, the leaching temperature can be 50°C, 55°C, 60°C, 65°C, 70°C, etc., and the leaching time can be 1h, 2h, 3h, 4h, etc. After the leaching is completed, solid-liquid separation can be performed by filtering or other means, and the filter residue is used as the iron ore concentrate (part) used for the next round of leaching, and the filtrate enters step S2.

[0053] It should be noted that by optimizing the raw materials and operating parameters of hydrochloric acid leaching, the leaching rate of iron ore concentrate can be higher than 95%.

[0054] S2, Iron powder reduction

[0055] The hydrochloric acid leaching solution and iron powder are mixed and reacted, and most of the Cu is removed by reduction substitution using the iron powder. After the reaction is completed, the solid-liquid separation is performed to obtain a reduction leaching solution.

[0056] In some embodiments, the reaction temperature of iron powder reduction is 50°C-70°C, the reaction time is 1h-3h, and the pH value of the solution at the reaction end point is controlled to be greater than 2.0. By regulating the temperature and time of iron powder reduction, impurities can be more fully replaced and the impurity removal effect can be improved. Specifically, the reaction temperature can be 50°C, 55°C, 60°C, 65°C, 70°C, etc., and the reaction time can be 1h, 2h, 3h, etc. After reaching the end point, stop adding iron powder, and use filtration or other means to separate solid and liquid, and use the obtained filter residue as the iron powder (part) for the next round of reduction.

[0057] It should be noted that the main impurity elements in the iron ore concentrate leaching solution are Al, Cr, Co, Ni, and Cu. The use of iron powder reduction can replace and remove most of the Cu and a small amount of Co, while raising the solution pH to 2-3, so that most of the Al 3+ Cr 3+ No H2S gas is produced during hydrolysis and sulfidation.

[0058] In some embodiments, during the reduction of iron powder, the ratio of the amount of iron powder added to the theoretical amount of iron powder is controlled to be (1.0-1.1):1, such as 1.00:1, 1.03:1, 1.05:1, 1.08:1, 1.10:1, etc. The amount of iron powder used is preferably slightly excessive so that impurities in the hydrochloric acid leachate can be more fully removed.

[0059] S3, sulfurization and impurity removal

[0060] The reduction leaching solution is mixed with sulfide to react, so that impurities such as Ni and Co are precipitated. After the reaction is completed, the solid and liquid are separated to obtain the reduced liquid. Taking advantage of the fact that the Ksp of NiS and CoS is smaller than that of FeS, sulfide is used for sulfidation and impurity removal, so that Ni and Co precipitation and recovery can be completed without losing Fe.

[0061] Sulfide removal refers to the difference in Ksp values ​​between different sulfides. By adding sulfides with higher solubility, the sulfides and Ni in the solution are separated. 2+ 、Co 2+ The plasma undergoes a replacement reaction to generate NiS, CoS and other precipitates with smaller Ksp, thereby achieving the purpose of purifying the solution and removing impurities.

[0062] In some embodiments, the sulfide is selected from at least one of sodium sulfide, ferrous sulfide, and manganese sulfide, and the sulfide can be any one or more of the above. The process of sulfidation and impurity removal includes: mixing a theoretical amount of sulfide with 5%-10% of the reducing leaching solution, separating the solid and liquid to obtain a slag, mixing the slag with the remaining reducing leaching solution, and then separating the solid and liquid to obtain a reduced liquid. First, all the required amount of sulfide is reacted with a small amount of ferrous chloride in the reducing leaching solution to cause iron and impurity elements to be deposited into the slag. Later, the slag is mixed with the remaining reducing leaching solution to avoid the introduction of impurities such as sodium into the reduced liquid. First, the theoretical amount of sulfide is mixed with 5%-10% of the reducing leaching solution to obtain crude ferrous sulfide, and then ferrous sulfide is used to replace and remove Ni. 2+ 、Co 2+ , Mn 2+ etc. impurity ions, reducing Na + etc. introduced.

[0063] Specifically, the theoretical amount refers to the amount of sulfide theoretically consumed by the nickel and cobalt elements in the total amount of the reduction leaching solution, that is, the theoretical amount of sulfide refers to the amount of sulfide required calculated based on the total content of Ni and Co in the reduction leaching solution (according to the measured Ni and Co concentrations, sulfides are added at a molar ratio of 1:1). The amount of reduction leaching solution added when preparing the slag accounts for 5%-10% of the total mass of the reduction leaching solution, which can be 5%, 6%, 7%, 8%, 9%, 10%, etc.

[0064] Further, the reaction time with 5%-10% reduction leaching solution is 30min-60min, such as 30min, 40min, 50min, 60min, etc. The reaction temperature with the remaining reduction leaching solution is 60℃-85℃, such as 60℃, 65℃, 70℃, 75℃, 80℃, 85℃, etc. The reaction time is 1h-2h, such as 1.0h, 1.5h, 2.0h, etc. After the reaction with the remaining reduction leaching solution is completed, it is filtered, the filter residue is NiCu enriched residue, and the filtrate is the reduced liquid.

[0065] In some embodiments, before the slag is mixed with the remaining reducing leaching solution for reaction, the slag is washed, filtered, and slurried to remove residual ions on the surface, and then water is added to slurry to facilitate transfer.

[0066] S4, evaporation crystallization

[0067] The reduced solution is heated and evaporated, and the solution is heated to boil and the solvent evaporates, so that the solution changes from an unsaturated state or a saturated state to a supersaturated state, and part of the water in the solution is removed. Then, the solution is cooled and crystallized, so that the solubility of the solute decreases, the supersaturation increases, and finally the supersaturated solute crystals are precipitated while the trace impurities remain in the solution, so as to achieve the effect of purifying the solute. After cooling and crystallizing, solid-liquid separation is performed to obtain crude ferrous chloride.

[0068] In some embodiments, the evaporation method is vacuum distillation, the operating pressure is controlled to be 0.005MPa-0.020MPa, and the evaporation ratio is controlled to be 15%-25%. The vacuum distillation method can improve the evaporation efficiency and quickly remove water from the solution. The operating pressure during the vacuum distillation process can be 0.005MPa, 0.010MPa, 0.015MPa, 0.020MPa, etc. The evaporation ratio refers to the ratio of the volume of water removed by evaporation to the original volume, and the evaporation ratio can be 15%, 20%, 25%, etc.

[0069] In some embodiments, the temperature is lowered to 30°C-40°C after evaporation, and the cooling crystallization is performed for 3h-5h. By regulating the temperature and time of the cooling crystallization, the ferrous chloride is more fully precipitated and the recovery rate is improved. Specifically, the temperature of the cooling crystallization can be 30°C, 35°C, 40°C, etc., and the time of the cooling crystallization can be 3h, 4h, 5h, etc.

[0070] Furthermore, after cooling and crystallization, solid-liquid separation can be performed by filtering or other means, and the obtained filtrate is used as the bottom liquid of the next cycle, and the obtained crystals are crude ferrous chloride.

[0071] S5. Washing

[0072] The crude ferrous chloride is washed with a 2 mol / L ferrous chloride solution, and solid-liquid separation is performed after the washing to obtain ferrous chloride crystals. The impurities on the surface of the crude ferrous chloride are further removed by washing to improve the purity of the product. The ferrous chloride solution used in the washing process is a nearly saturated ferrous chloride solution to further reduce the loss of raw materials and improve the recovery rate.

[0073] In some embodiments, the pulp washing process includes: mixing crude ferrous chloride and a nearly saturated ferrous chloride solution for pulp washing, and then using filtration or other means to separate the solid and liquid to obtain a filtrate and a filter residue, and the obtained filter residue is a high-purity ferrous chloride crystal. The obtained 1%-5% filtrate is used as the ferrous chloride bottom liquid required for the hydrochloric acid leaching process in step S1, and the remaining 95%-99% of the filtrate is supplemented with water to 100% of the original volume and used as a nearly saturated ferrous chloride solution for the next cycle of pulp washing, so as to fully recover the filtrate and reduce the process cost.

[0074] In some embodiments, during the pulping process, the volume of ferrous chloride solution corresponding to 1 kg of crude ferrous chloride is controlled to be 0.5 L-1.0 L, such as 0.5 L, 0.6 L, 0.7 L, 0.8 L, 0.9 L, 1.0 L, etc. The pulping time is controlled to be 10 min-30 min, such as 10 min, 20 min, 30 min, etc.

[0075] S6. Calcination

[0076] The ferrous chloride crystals are roasted in a water-containing air atmosphere to obtain high-purity iron red and high-purity iron yellow. By controlling the roasting temperature, high-purity iron yellow and iron red can be selectively prepared using the same materials and process.

[0077] It should be noted that, currently, the decomposition of ferrous chloride to produce iron oxide red in industry mostly adopts the spray method, which requires the use of special spray equipment. According to the mechanism that the thermal decomposition of ferrous chloride requires the participation of water, the present invention introduces humidified air into the tubular furnace for roasting, and can efficiently convert ferrous chloride into iron oxide red without using special equipment, thereby reducing equipment investment.

[0078] Specifically, the high-purity iron oxide prepared in the embodiment of the present invention is a first-class iron oxide product that meets the YHT1 index of the GB / T24244-2009 standard.

[0079] In some embodiments, the ferrous chloride crystals are dried and then calcined, the calcination temperature is controlled to be 300°C-500°C, and the calcination time is 1h-5h. The calcination temperature and time are adjusted to allow the ferrous chloride to fully react to obtain high-purity iron red or high-purity iron yellow. Specifically, the calcination temperature can be 300°C, 350°C, 400°C, 450°C, 500°C, etc., and the calcination time can be 1h, 2h, 3h, 4h, 5h, etc. The powder obtained after calcination is washed and dried to further improve the purity of the product. The number of washings can be 2 times, but is not limited thereto.

[0080] Furthermore, the waste gas generated during the roasting process is introduced into a spray device to obtain a hydrochloric acid solution with a concentration of 1 mol / L-3 mol / L, so as to further recover the generated gas. The roasting device used can be an air-blowing tube furnace, and the air blowing rate is controlled to be 2L / min-3L / min, such as 2.0L / min, 2.5L / min, 3.0L / min, etc.

[0081] The features and performance of the present invention are further described in detail below in conjunction with the embodiments.

[0082] It should be noted that the main metal components of the raw material iron ore concentrate in the following embodiments are as follows: Fe (55% to 65%), Co (0.01% to 0.1%), Ni (1% to 2.5%), Cu (1% to 2%), Al (0.0% to 0.5%). In addition, the iron ore concentrate also contains a small amount of silicon, zinc, chromium, etc., and can be regarded as a low-quality iron red rich in impurities, in which the iron element mainly exists in the form of Fe2O3, and Fe presents a +3 valence.

[0083] Example 1

[0084] This embodiment provides a method for recovering iron ore concentrate, comprising the following steps:

[0085] (1) Add 0.55 L of 36% (mass fraction, the same below) concentrated hydrochloric acid to 1.55 L of saturated ferrous chloride solution at room temperature, add 200 g of iron concentrate to the leaching bottom solution at 60°C and 300 rpm, react for 2 h, and filter to obtain a hydrochloric acid leaching solution.

[0086] (2) Slowly add 70 g of iron powder to the hydrochloric acid leaching solution, control the temperature to 70-80° C., react for 2 h, and filter to obtain a reduced leaching solution.

[0087] (3) Take 5% reduction leaching solution, add theoretical amount of industrial sodium sulfide thereto, react at room temperature for 30 minutes, filter, wash the filter residue twice, add the obtained FeS to the remaining 95% iron powder reduction solution, react at 70°C for 1 hour, filter, and obtain the reduction solution.

[0088] (4) The reduced solution obtained by filtration was placed in a rotary evaporator to evaporate 400 mL of water. The evaporated solution was placed in a beaker and cooled and crystallized at 60 rpm. After the temperature dropped to room temperature, it was filtered. After the ferrous chloride crystals were washed twice, pure ferrous chloride crystals were filtered and dried at 60°C.

[0089] (5) Grind the dried ferrous chloride crystals, sieve the ground powder (200 mesh), take 30g of the sieve under the sieve and spread it in a corundum crucible (thickness of about 0.5cm), and place it in a tube furnace for roasting for 2h. The roasting temperature is 350℃. During the roasting process, air containing water mist is continuously introduced, and the air flow rate is controlled to be 2L / min. After cooling in the furnace, the powder is washed and dried to obtain high-purity iron yellow or high-purity iron red, which is weighed and bagged, and the corresponding quality analysis is performed.

[0090] Example 2

[0091] This embodiment provides a method for recovering iron ore concentrate, comprising the following steps:

[0092] (1) Add 0.55 L of 36% concentrated hydrochloric acid to 1.55 L of saturated ferrous chloride solution at room temperature, add 200 g of iron concentrate to the leaching bottom solution at 60°C and 300 rpm, react for 2 h, and filter to obtain a hydrochloric acid leaching solution.

[0093] (2) Slowly add 70 g of iron powder to the hydrochloric acid leaching solution, control the temperature to 70-80° C., react for 2 h, and filter to obtain a reduced leaching solution.

[0094] (3) Take 5% of the reduced leaching solution, add a theoretical amount of industrial sodium sulfide thereto, react for 30 minutes, filter, wash the filter residue twice, add the obtained FeS to the remaining 95% iron powder reduced solution, react at 70°C for 1 hour, filter, and obtain a reduced solution.

[0095] (4) The reduced solution obtained by filtration was placed in a rotary evaporator to evaporate 400 mL of water. The evaporated solution was placed in a beaker and cooled and crystallized at 60 rpm. After the temperature dropped to room temperature, it was filtered. After the ferrous chloride crystals were washed twice, pure ferrous chloride crystals were filtered and dried at 60°C.

[0096] (5) Grind the dried ferrous chloride crystals, sieve the ground powder (200 mesh), take 30g of the sieve under the sieve and spread it in a corundum crucible (thickness of about 0.5cm), and place it in a tube furnace for roasting for 2h at a roasting temperature of 400°C. During the roasting process, air containing water mist is continuously introduced, and the air flow rate is controlled at 2L / min. After cooling in the furnace, the powder is washed and dried to obtain high-purity iron yellow or high-purity iron red, which is weighed and bagged, and the corresponding quality analysis is performed.

[0097] The only difference between Example 2 and Example 1 is that the calcination temperature in step (5) is 400°C.

[0098] Example 3

[0099] This embodiment provides a method for recovering iron ore concentrate, comprising the following steps:

[0100] (1) Add 0.55 L of 36% concentrated hydrochloric acid to 1.55 L of saturated ferrous chloride solution at room temperature, add 200 g of iron concentrate to the leaching bottom solution at 60°C and 300 rpm, react for 2 h, and filter to obtain a hydrochloric acid leaching solution.

[0101] (2) Slowly add 70 g of iron powder to the hydrochloric acid leaching solution, control the temperature to 70-80° C., react for 2 h, and filter to obtain a reduced leaching solution.

[0102] (3) Take 5% iron powder reduction leaching solution, add the theoretical amount of industrial sodium sulfide thereto, react for 30 minutes, filter, wash the filter residue twice, add the obtained FeS to the remaining 95% iron powder reduction solution, react at 70°C for 1 hour, filter, and obtain the reduction solution.

[0103] (4) The reduced solution obtained by filtration was placed in a rotary evaporator to evaporate 400 mL of water. The evaporated solution was placed in a beaker and cooled and crystallized at 60 rpm. After the temperature dropped to room temperature, it was filtered. After the ferrous chloride crystals were rinsed twice, pure ferrous chloride crystals were filtered and dried at 60°C.

[0104] (5) Grind the dried ferrous chloride crystals, sieve the ground powder (200 mesh), take 30g of the sieve under the sieve and spread it in a corundum crucible (thickness of about 0.5cm), and place it in a tube furnace for roasting for 2h at a roasting temperature of 400°C. During the roasting process, air containing water mist is continuously introduced, and the air flow rate is controlled at 2L / min. After cooling in the furnace, the powder is washed and dried to obtain high-purity iron yellow or high-purity iron red, which is weighed and bagged, and the corresponding quality analysis is performed.

[0105] The only difference between Example 3 and Example 2 is that pulp washing is replaced by rinsing in step (4).

[0106] Comparative Example 1

[0107] (1) Add 0.55 L of 36% concentrated hydrochloric acid to 1.55 L of saturated ferrous chloride solution at room temperature, add 200 g of iron concentrate to the leaching bottom solution at 60°C and 300 rpm, react for 2 h, and filter to obtain a hydrochloric acid leaching solution.

[0108] (2) Slowly add 70 g of iron powder to the hydrochloric acid leaching solution, control the temperature not higher than 80° C., react for 2 h, filter, and obtain a reduced leaching solution.

[0109] (3) Take 5% iron powder reduction leaching solution, add the theoretical amount of industrial sodium sulfide thereto, react for 30 minutes, filter, wash the filter residue twice, add the obtained FeS to the remaining 95% iron powder reduction solution, react at 70°C for 1 hour, filter, and obtain the reduction solution.

[0110] (4) The reduced solution obtained by filtration was placed in a rotary evaporator to evaporate 400 mL of water. The evaporated solution was placed in a beaker and cooled and crystallized at 60 rpm. After the temperature dropped to room temperature, it was filtered. After the ferrous chloride crystals were washed twice, pure ferrous chloride crystals were filtered and dried at 60°C.

[0111] (5) Grind the dried ferrous chloride crystals, sieve the ground powder (200 mesh), take 30g of the sieve under the sieve and spread it in a corundum crucible (thickness of about 0.5cm), and place it in a tube furnace for roasting for 2h. The roasting temperature is 500℃. During the roasting process, air containing water mist is continuously introduced, and the air flow rate is controlled at 2L / min. After cooling in the furnace, the powder is washed and dried to obtain high-purity iron yellow or high-purity iron red, which is weighed and bagged, and the corresponding quality analysis is performed.

[0112] The only difference between Comparative Example 1 and Example 1 is that the calcination temperature in step (5) is 500°C.

[0113] Comparative Example 2

[0114] (1) Add 0.55 L of 36% concentrated hydrochloric acid to 1.55 L of saturated ferrous chloride solution at room temperature, add 200 g of iron concentrate to the leaching bottom solution at 60°C and 300 rpm, react for 2 h, and filter to obtain a hydrochloric acid leaching solution.

[0115] (2) Slowly add 70 g of iron powder to the hydrochloric acid leaching solution, control the temperature not higher than 80° C., react for 2 h, filter, and obtain a reduced leaching solution.

[0116] (3) Take 5% of the reduced leaching solution, add a theoretical amount of industrial sodium sulfide thereto, react for 30 minutes, filter, wash the filter residue twice, add the obtained FeS to the remaining 95% iron powder reduced solution, react at 70°C for 1 hour, filter, and obtain a reduced solution.

[0117] (4) The reduced solution obtained by filtration was placed in a rotary evaporator to evaporate 400 mL of water. The evaporated solution was placed in a beaker and cooled and crystallized at 60 rpm. After the temperature dropped to room temperature, it was filtered. After the ferrous chloride crystals were washed twice, pure ferrous chloride crystals were filtered and dried at 60°C.

[0118] (5) Grind the dried ferrous chloride crystals, sieve the ground powder (200 mesh), take 30 g of the sieve under the sieve and spread it in a corundum crucible (thickness of about 0.5 cm), and place it in a tube furnace for roasting for 2 hours at a roasting temperature of 400°C. During the roasting process, water mist-containing air is continuously introduced at an air flow rate of 2 L / min. After cooling in the furnace, the powder is weighed and bagged without washing, and the corresponding quality analysis is performed.

[0119] The only difference between Comparative Example 2 and Example 2 is that step (5) does not involve washing and drying at the end.

[0120] Comparative Example 3

[0121] (1) Add 0.55 L of 36% concentrated hydrochloric acid to 1.55 L of saturated ferrous chloride solution at room temperature, add 200 g of iron concentrate to the leaching bottom solution at 60°C and 300 rpm, react for 2 h, and filter to obtain a hydrochloric acid leaching solution.

[0122] (2) Slowly add 70 g of iron powder to the hydrochloric acid leaching solution, control the temperature not higher than 80° C., react for 2 h, adjust the pH to 4.5 using liquid alkali, react for 1 h, filter, and obtain a pH-adjusted impurity-removing liquid.

[0123] (3) The reduced leaching solution obtained by filtration was placed in a rotary evaporator to evaporate 400 mL of water. The evaporated solution was placed in a beaker and cooled and crystallized at 60 rpm. After the temperature dropped to room temperature, it was filtered. After the ferrous chloride crystals were washed twice, pure ferrous chloride crystals were filtered and dried at 60°C.

[0124] (4) Grind the dried ferrous chloride crystals, sieve the ground powder (200 mesh), take 30g of the sieve under the sieve and spread it in a corundum crucible (thickness of about 0.5cm), and place it in a tube furnace for roasting for 2h. The roasting temperature is 400℃. During the roasting process, air containing water mist is continuously introduced, and the air flow rate is controlled at 2L / min. After cooling in the furnace, the powder is washed and dried to obtain high-purity iron yellow or high-purity iron red, which is weighed and bagged, and the corresponding quality analysis is performed.

[0125] Comparative Example 4

[0126] The only difference from Example 1 is that the specific steps of step (3) are different, which are as follows: the reduced leaching solution obtained by filtration is placed in a rotary evaporator, 400 mL of water is evaporated, the evaporated solution is placed in a beaker and cooled and crystallized at 60 rpm, and after the temperature drops to room temperature, it is filtered, and the ferrous chloride crystals are not washed and directly dried at 60°C.

[0127] Comparative Example 5

[0128] The difference from Example 1 is that the specific steps of step (5) are different, which are as follows: grind the dried ferrous chloride crystals, sieve the ground powder (200 mesh), take 30g of the sieve under the sieve and spread it in a corundum crucible (thickness of about 0.5cm), place it in a tube furnace and roast it for 2h, the roasting temperature is 250℃, and the air containing water mist is continuously introduced during the roasting process, and the air flow rate is controlled to be 2L / min. After cooling in the furnace, the powder is washed and dried to obtain high-purity iron yellow or high-purity iron red, weighed and bagged, and the corresponding quality analysis is performed.

[0129] Test Example 1

[0130] The impurity removal effects and product qualities of the test examples and comparative examples are shown in Tables 1 to 3.

[0131] Table 1 Element composition analysis of the sulfurized impurity-removing liquid obtained in Example 1 and the pH-adjusted impurity-removing liquid of Comparative Example 3 (g / L)

[0132] Group Fe Co Ni Cu Mn Cr Pb S Comparative Example 3 pH Adjustment and Impurity Removal Liquid 119.7 0.0182 1.491 0.0008 0.2714 0.0095 0.0107 0.3256 Example 1: Sulfurization and impurity removal liquid 243.9 0.0301 1.600 0.0022 0.3847 0.0677 0.0095 4.783

[0133] Table 1 is the elemental composition analysis of the sulfidation impurity removal liquid of Example 1 and the pH-adjusted impurity removal liquid of Comparative Example 3. The results show that if the impurities are removed by adjusting the pH to a sufficient depth, the loss of the main metal element Fe is huge, while sulfidation impurity removal can avoid the precipitation loss of Fe.

[0134] Table 2. Purity analysis of pure ferrous chloride in Example 2 and Example 3 (%)

[0135] Group Fe Co Ni Cu Mn Cr Ca S Example 2 28.01 0.0027 0.0031 0.0007 0.0255 0.0033 0.0004 Not detected Example 3 27.78 0.0015 0.0006 0.0001 0.0298 0.0018 0.1060 0.0241

[0136] Table 2 compares the purity of ferrous chloride crystals obtained by pulp washing (Example 2) and elution washing (Example 3). It can be seen that the ferrous chloride crystals prepared by pulp washing are significantly better than those prepared by elution washing in terms of S and Ca indicators.

[0137] Table 3. Analysis of iron red components in Example 2 and Comparative Example 2

[0138]

[0139] Table 3 compares the result of elemental analysis when the iron oxide red is washed with water and not washed after the roasting. It can be seen that the decomposition rate of FeCl2 in the roasting process does not reach 100%, and residual FeCl2 needs to be removed by washing. After the roasting, the powder can be dried to obtain the final product. In Table 3, Example 2 shows the result of elemental analysis of the final product. The iron oxide red product obtained in this experiment, Fe2O3 mass fraction is 99.414%, which meets the requirement that the YHT1 iron oxide red Fe2O3 mass fraction is greater than 99.40% in GB / T24244-2009 "Iron Oxide for Ferrites". In addition, the foreign ion content all reaches the requirement of GB / T24244-2009 standard YHT1 iron oxide red.

[0140] Test Example 2

[0141] The chromaticity of the products prepared in Examples 1-2 and Comparative Example 1 was tested. The results are as follows: Figure 2 shown.

[0142] from Figure 2 It can be seen that the difference in chromaticity between the embodiment and the comparative example shows that the calcination temperature has a great influence on the hue of iron oxide red, and bright red iron oxide red can only be prepared at a temperature close to 400°C.

[0143] The above are only preferred embodiments of the present invention and are not intended to limit the present invention. For those skilled in the art, the present invention may have various modifications and variations. Any modification, equivalent replacement, improvement, etc. made within the spirit and principle of the present invention shall be included in the protection scope of the present invention.

Claims

1. A method for recovering iron ore concentrate, characterized in that: include: Hydrochloric acid leaching: The iron ore concentrate is mixed with a leaching solution containing hydrochloric acid for reaction, and then the solid and liquid are separated to obtain a hydrochloric acid leaching solution; Iron powder reduction: mixing the hydrochloric acid leaching solution and iron powder for reaction, and then performing solid-liquid separation to obtain a reduced leaching solution; Sulfidation and impurity removal: mixing the reduced leaching solution with sulfide for reaction, and then separating the solid and liquid to obtain the reduced liquid; Evaporation crystallization: heating the reduced liquid to evaporate, then cooling to crystallize, and separating the solid from the liquid to obtain crude ferrous chloride; Washing: washing the crude ferrous chloride with a ferrous chloride solution, and separating the solid from the liquid to obtain ferrous chloride crystals; Calcination: calcining the ferrous chloride crystals, and washing the powder obtained after calcination with water.

2. The method according to claim 1, characterized in that The leaching solution used in the hydrochloric acid leaching process is obtained by mixing hydrochloric acid and ferrous chloride base solution; Optionally, the preparation process of the leaching solution includes: mixing concentrated hydrochloric acid with a mass fraction of 30%-40% and a ferrous chloride base solution with a concentration of 2.8mol / L-3.2mol / L, wherein the volume ratio of the concentrated hydrochloric acid to the ferrous chloride base solution is 1:(2.5-3.0); Optionally, the mass ratio of the iron ore concentrate to the leaching solution is 1:(8-12).

3. The method according to claim 1 or 2, characterized in that: The leaching temperature of the hydrochloric acid leaching is 50° C.-70° C., and the leaching time is 1 h-4 h.

4. The method according to claim 1, characterized in that: During the reduction of the iron powder, the ratio of the amount of iron powder added to the theoretical amount of iron powder is controlled to be (1.0-1.1):1; Optionally, the reaction temperature of the iron powder reduction is 50° C.-70° C., the reaction time is 1 h-3 h, and the pH value of the solution at the reaction endpoint is controlled to be greater than 2.

0.

5. The method according to claim 1, characterized in that The sulfidation impurity removal process comprises: mixing a theoretical amount of sulfide with 5%-10% of the reduction leaching solution, separating the solid and the liquid to obtain a slag material, mixing the slag material with the remaining reduction leaching solution, and then separating the solid and the liquid to obtain the reduced liquid; wherein the theoretical amount refers to the theoretical amount of sulfide required for the nickel and cobalt elements in the total amount of the reduction leaching solution; Optionally, the sulfide is selected from at least one of sodium sulfide, ferrous sulfide, and manganese sulfide; Optionally, the reaction time with 5%-10% of the reducing leaching solution is 30min-60min; Optionally, the reaction temperature with the remaining reduction leaching solution is 60° C.-85° C., and the reaction time is 1 h-2 h; Optionally, the slag is washed and slurried before being mixed with the remaining reducing leaching solution for reaction.

6. The method according to claim 1, characterized in that During the evaporation crystallization process, the evaporation ratio is controlled to be 15%-25%; Optionally, the evaporation method is reduced pressure distillation, and the operating pressure is controlled to be 0.005MPa-0.020MPa.

7. The method according to claim 1 or 6, characterized in that: After evaporation, cool to 30-40°C and cool for crystallization for 3h-5h; Optionally, after cooling and crystallization, the solid and liquid are separated, and the obtained filtrate is used as the bottom liquid of the next cycle.

8. The method according to claim 1, characterized in that The crude ferrous chloride is pulp-washed with a ferrous chloride solution having a concentration of 1.5-2.5 mol / L; Optionally, during the pulp washing process, the volume of the ferrous chloride solution corresponding to 1 kg of the crude ferrous chloride is controlled to be 0.5 L-1.0 L; Optionally, the pulping time is controlled to be 10 min-30 min.

9. The method according to claim 1 or 8, characterized in that: The pulping process includes: mixing the crude ferrous chloride and the ferrous chloride solution for pulping, then separating the solid and the liquid, using the obtained 1%-5% filtrate as the ferrous chloride bottom liquid required for the hydrochloric acid leaching process, and replenishing the remaining filtrate with water as the ferrous chloride solution for the next cycle of pulping.

10. The method according to claim 1, characterized in that The calcination is carried out in a water-containing air atmosphere, the calcination temperature is controlled to be 300°C-500°C, and the calcination time is 1h-5h; Optionally, the ferrous chloride crystals are dried and then calcined; Optionally, the waste gas generated during the roasting process is introduced into a spray device to obtain a hydrochloric acid solution with a concentration of 1 mol / L to 3 mol / L; Optionally, the roasting device used is an air-blast tube furnace, and the air blowing rate is controlled to be 2L / min-3L / min; The method further comprises: washing the powder obtained after calcination with water and then drying it.

Images