Method for preparing magnetite concentrate and tin-rich concentrate from tin-iron tailings

Abstract

The invention discloses a method for preparing magnetite concentrate and tin-rich concentrate from tin-iron tailings, and belongs to the technical field of tailing resource utilization. The method comprises the following steps: uniformly mixing the tin-iron tailings and calcium carbonate minerals, carrying out agglomeration, and carrying out reduction roasting on the obtained agglomerates in a CO and CO2 mixed atmosphere to obtain a roasted product containing calcium stannate and magnetite; and after the roasted product is subjected to ball milling and size mixing, magnetite concentrate recovery through magnetic separation, gangue removal through gravity separation and calcium stannate concentrate recovery through flotation are sequentially carried out, and by means of the method, efficient conversion of the tin-iron tailings can be achieved at the low temperature, efficient enrichment and separation of tin minerals and iron minerals can be achieved, and the yield of the tin-iron tailings is improved. And high-grade calcium stannate concentrate products and magnetite products are obtained, and efficient, clean and large-scale utilization of tin-iron tailing resources is truly achieved.

Description

Technical Field

[0001] The invention relates to a method for processing tin-iron tailings, in particular to a method for preparing magnetite concentrate and tin-rich concentrate from the tin-iron tailings, and belongs to the technical field of comprehensive utilization of nonferrous metal resources. Background Art

[0002] Tin-iron tailings are a complex and difficult-to-process ore with relatively abundant reserves but low levels of development and utilization. These ores typically contain 30% to 55% iron and 0.1% to 0.5% tin, with some high-grade ores containing tin levels exceeding 0.8%. Because they are rich in iron and contain the economically valuable element tin, they possess significant potential for resource utilization. However, due to their complex mineral composition and difficulty in processing, they have long remained underdeveloped.

[0003] Tin in tin-iron tailings primarily exists as cassiterite (SnO2), present in an extremely finely interbedded state. It often occurs as a single particle or closely intergrows with iron oxides (such as hematite and limonite), resulting in a dense structure that makes it difficult to effectively separate them using conventional physical beneficiation methods. This complex mineral intergrowth leads to a blurred interface between tin and iron, causing co-floating or co-settling during the beneficiation process. This results in poor tin-iron separation selectivity, limited separation performance, and severely impacts comprehensive resource recovery.

[0004] Currently, pyrometallurgical methods are primarily used to separate tin and iron from tin-iron tailings. Typical methods include sulfide volatilization, chlorination roasting, and weak reduction volatilization. The sulfide volatilization method effectively removes tin by generating SnS that volatilizes at high temperatures, but the operating temperature typically requires 1180°C to 1300°C and is accompanied by large amounts of SO2 gas emissions, placing a significant burden on equipment and the environment. The chlorination roasting method relies on a chlorinating agent to react with tin and iron to generate volatile chlorides, allowing tin to be volatilized and recovered at around 1000°C. However, this method presents challenges such as severe equipment corrosion and chlorine contamination. The weak reduction volatilization method utilizes the high vapor pressure of SnO at high temperatures to achieve volatilization, but requires stringent atmosphere and temperature control and is not conducive to iron recovery.

[0005] Currently, patented technologies utilize low-temperature reduction roasting combined with magnetic separation and leaching to comprehensively utilize tin and iron ore. For example, Chinese patent (CN104152671B) proposes using sodium carbonate as an additive to roast tin and iron ore in a weakly reducing atmosphere of CO / CO₂. Tin enters the solution as sodium stannate, while iron is separated as magnetite. After magnetic separation and filtration, magnetite concentrate and sodium stannate solution are obtained, respectively. This method has a low roasting temperature (850°C to 945°C), low energy consumption, excellent tin-iron separation, and is environmentally friendly. However, the recovery of residual tin in the tailings requires a wet process, which involves multiple steps and complex operations. The tin-rich product is unstable in form, making subsequent resource utilization difficult. Summary of the Invention

[0006] In view of the technical problems existing in the prior art of tin-iron tailings treatment processes, such as high roasting temperature, complex process, incomplete tin-iron separation, and low utilization rate of tin resources in the tailings, the present invention aims to provide a method for preparing magnetite concentrate and tin-rich concentrate from tin-iron tailings. The method can not only achieve efficient conversion of tin-iron tailings at a relatively low temperature, but also achieve efficient enrichment and separation of tin minerals and iron minerals, and obtain high-grade calcium stannate products and magnetite products, thereby truly realizing efficient, clean and large-scale utilization of tin-iron tailings resources.

[0007] In order to achieve the above technical objectives, the present invention provides a method for preparing magnetite concentrate and tin-rich concentrate from tin-iron tailings. The method comprises the following steps: mixing the tin-iron tailings with calcium carbonate minerals to form agglomerates, and then subjecting the resulting agglomerates to reduction roasting in a mixed atmosphere of CO and CO2 to obtain a roasted product containing calcium stannate and magnetite; and the roasted product is subjected to ball milling and slurry mixing, and then subjected to magnetic separation to recover magnetite concentrate, gravity separation to remove gangue, and flotation to recover calcium stannate concentrate.

[0008] The key to the technical solution of the present invention is to first use calcium carbonate as an additive to achieve low-temperature conversion of tin-iron tailings, convert the tin-iron tailings into calcium stannate minerals and magnetite minerals, and then use magnetic separation, gravity separation and flotation processes to achieve separation of calcium stannate minerals and magnetite minerals, thereby obtaining high-grade calcium stannate concentrate and magnetite concentrate. More specifically, the present invention uses calcium carbonate as an additive, and the tin-iron tailings are reduced and roasted in a mixed atmosphere of CO and CO2. The reduction roasting process can convert the high-valent iron oxides in the tin-iron tailings into magnetite, and activate cassiterite, while decomposing calcium carbonate at high temperature to form active calcium oxide, and active calcium oxide plays an important role in the roasting process. On the one hand, it can reduce the roasting temperature of the tin-iron tailings as a flux, and on the other hand, it can capture and convert cassiterite into calcium stannate minerals as a conversion agent of tin oxide. The tin-iron is transformed into magnetite minerals and calcium stannate minerals by reduction roasting, providing favorable conditions for the subsequent mineral separation process. In the mineral separation process, it is preferred to utilize the strong magnetic properties of magnetite and first separate it from calcium stannate minerals through magnetic separation to obtain high-grade magnetite concentrate. However, in the magnetic separation tailings, the grade of calcium stannate minerals is relatively low. The present invention adopts a method of combining gravity separation and tailings discarding with flotation enrichment to achieve the recovery of calcium stannate minerals. After centrifugal gravity separation, most gangue minerals can be efficiently removed, the initial enrichment of calcium stannate minerals is achieved, and the subsequent flotation selection grade and separation boundary are improved. On this basis, further tin-enriched concentrate of calcium stannate minerals is achieved through flotation to obtain high-grade calcium stannate minerals.

[0009] As a preferred solution, the mass ratio of the tin-iron tailings to the calcium carbonate mineral is 1:0.2-0.4. The ratio of the tin-iron tailings to the calcium carbonate mineral is preferably controlled within a preferred range. If the calcium carbonate ratio is too low, the cassiterite conversion rate will be low, the recovery will be incomplete, and the required roasting temperature will be high. If the calcium carbonate ratio is too high, some iron minerals will be converted into minerals such as calcium ferrite that cannot be recovered by magnetic separation, resulting in a low iron recovery rate.

[0010] As a preferred embodiment, the tin-iron tailings contain 30% to 55% TFe by mass and 0.2% to 0.8% Sn by mass. Tin-iron tailings are characterized by high iron grade and low tin grade. The tin primarily exists in the form of cassiterite (SnO2), which is embedded in extremely fine particles, often as single particles or closely associated with iron oxides (such as hematite and limonite), resulting in a dense structure.

[0011] As a preferred embodiment, the calcium carbonate mineral comprises at least one of industrial-grade calcium carbonate, limestone, and calcite, with a CaCO3 mass content of 95% or greater. The calcium carbonate mineral can be derived from industrial by-product calcium carbonate, mineral processing tailings, or industrial analytical grade reagents. The addition of the calcium carbonate mineral serves, on the one hand, to adjust the mineral alkalinity, and, on the other hand, to enhance the activation reaction behavior of cassiterite.

[0012] The tin-iron tailings and calcium carbonate minerals of the present invention are ground to a particle size of -0.1 mm, wherein fineness control is helpful to improve the reaction efficiency and separation limit between minerals.

[0013] As a preferred solution, the conditions for the reduction roasting are: the atmosphere is a mixed atmosphere of CO and CO2 with a CO volume concentration of 5% to 17%, the temperature is 850°C to 950°C, and the time is 30 to 60 minutes. Under the coordinated control of the preferred roasting temperature and atmospheric conditions, the reduction of high-valent iron minerals can be achieved, such as reducing and converting minerals such as hematite and limonite into magnetite minerals under a weak reducing atmosphere, while the decomposition of calcium carbonate minerals can be achieved, and the conversion of active calcium oxide and cassiterite into calcium stannate minerals can be achieved. If the roasting temperature is too low, the conversion of tin-iron tailings will be incomplete. If the roasting temperature is too high, the iron minerals will be converted into new phases such as calcium ferrite, reducing the recovery rate of iron. In addition, the reducing atmosphere affects the magnetization of high-valent iron minerals and the activation of cassiterite. If the CO volume concentration is too low, the magnetization of high-valent iron minerals will be incomplete, and the cassiterite will not be fully activated, reducing its recovery rate. If the CO volume concentration is too high, it is easy for some cassiterite to be reduced, which is not conducive to the recovery of tin. Through reduction roasting, cassiterite can be converted into calcium stannate minerals that are easy to float, while hematite and limonite are reduced to magnetite, thereby achieving differentiation of the mineral properties of the two.

[0014] As a preferred solution, the mass fraction of the -0.074 mm particle size is controlled to be above 90% during the ball milling process. Grinding the ore to an appropriate particle size facilitates subsequent magnetic separation, gravity separation, and flotation separation of the minerals.

[0015] As a preferred solution, the magnetic field strength used in the magnetic separation process is 800-1200 Gs. Under the preferred magnetic separation conditions, the magnetic product magnetite concentrate can be effectively separated, which has high iron grade and low tin content and is suitable for blast furnace smelting requirements.

[0016] As a preferred solution, a high-frequency centrifugal gravity separation method is used in the gravity separation process. The frequency of the high-frequency centrifugal gravity separation is 15-20 Hz, and the flow rate of the backwash water is controlled at 7-10 LPM. The non-magnetic minerals after magnetic separation contain calcium stannate minerals mainly composed of calcium stannate and a small amount of unreacted cassiterite, gangue, etc., which are subjected to preliminary density classification using a high-frequency centrifugal gravity separation process. Under the preferred gravity separation conditions, the efficient removal of gangue minerals and the enrichment of tin-containing minerals can be achieved, the subsequent flotation selection grade and separation boundary can be improved, and the tin enrichment multiple can be increased. Under these conditions, gangue inclusions such as quartz and calcite within the particle size range of 0.045-0.25 mm can be stably removed, and the tin grade of the gravity-separated mineral can be increased to above 1.7%.

[0017] As a preferred solution, during the flotation process, the pulp mass concentration is adjusted to 35-40%, and the pH is adjusted to 9-10. The preferred pulp concentration and pH are conducive to maintaining the stabilization of the surface charge properties of calcium stannate and the adsorption activation of the collector.

[0018] As a preferred embodiment, the flotation process utilizes a calcium stannate mineral collector composed of a fatty acid and a hydroxamic acid collector in a mass ratio of 2-4:1. The fatty acid and hydroxamic acid collectors are combined in an appropriate ratio. The fatty acid (such as oleic acid or its sodium salt) provides basic hydrophobic adsorption capacity, while the hydroxamic acid (such as sodium phenylhydroxamate) enhances the selective complexation of tin-containing phases such as calcium stannate. The synergistic effect of these two collectors significantly improves tin recovery efficiency and flotation concentrate quality. Ultimately, a high-grade, low-tin magnetite concentrate and a high-tin-grade, tin-rich concentrate are obtained, significantly increasing overall resource recovery and optimizing economic and environmental performance. In the flotation system, the fatty acid collector establishes initial hydrophobicity, while the hydroxamic acid enhances collection capacity and selectivity through chelation, effectively suppressing the co-floatation of associated iron and calcium impurities. This allows for highly selective tin enrichment, ultimately yielding a tin-rich concentrate with a tin grade of 3%-5%. The flotation process includes a roughing process, a cleaning process and a scavenging process.

[0019] During the agglomeration process of the tin-iron tailings and calcium carbonate minerals of the present invention, 8% to 15% water and 1% to 3% binder (such as bentonite or slaked lime) are added to form granules or briquettes. Agglomerates of 8 to 10 mm are produced, which facilitates uniform gas heat transfer and reduction reaction.

[0020] Compared with the prior art, the technical solution of the present invention brings the following beneficial technical effects:

[0021] (1) Although the traditional sulfide volatilization method can achieve high-temperature volatilization and recovery of tin, most of the iron is lost with the slag, making it impossible to obtain high-grade iron concentrate. Conventional flotation or gravity separation methods do not completely separate tin and iron, and the tin content in the magnetite is high, making it difficult to meet smelting requirements. The present invention uses controlled reduction roasting in a CO / CO2 atmosphere to effectively decouple cassiterite and iron oxide in structure. The tin and iron content of the magnetite obtained by magnetic separation after roasting fully meets the blast furnace raw material standards. At the same time, the tin-rich concentrate obtained by flotation has a tin grade of up to 3% to 5%, which is convenient for subsequent tin extraction and refining.

[0022] (2) The present invention takes a specific process as the main line. Compared with complex processes such as multi-stage wet leaching + multi-stage extraction of tin, the process links are clear and the parameter control range is wide (such as CO concentration, roasting temperature). It does not rely on highly corrosive chemicals or complex reagent systems. It has high industrial operability and can adapt to the strong volatility of different tin-iron tailings raw materials.

[0023] (3) Compared with traditional treatment methods, the present invention can not only effectively extract high-grade magnetite, but also achieve deep enrichment of non-magnetic tin resources, thereby minimizing the loss rate of tin in tailings. BRIEF DESCRIPTION OF THE DRAWINGS

[0024] Figure 1 The present invention is a process flow chart for preparing magnetite concentrate and tin-rich concentrate from tin-iron tailings. DETAILED DESCRIPTION

[0025] The following examples are intended to further illustrate the present invention, but are not intended to limit the scope of protection of the claims of the present invention.

[0026] In the following specific examples and comparative examples, the tin-iron tailings and calcium carbonate were ground to a particle size of -0.1 mm, with the proportion being 100%.

[0027] Comparative Example 1

[0028] The only difference between this comparative example and Example 1 is that no calcium carbonate additive is added, and other operations and parameters are the same as those in Example 1.

[0029] The resulting magnetite concentrate had a TFe content of 58.6%, a Sn content of 0.26%, and an iron recovery of 78.9%. Sn was not significantly enriched in the non-magnetic material, with a Sn content of 0.59%. These results indicate that the lack of calcium carbonate activation resulted in a lack of cassiterite activation, with tin primarily contained within the magnetic product, resulting in suboptimal tin-iron separation.

[0030] Comparative Example 2

[0031] The only difference between this comparative example and Example 2 is that the calcination temperature is 750° C., and the other operations and parameters are the same as those in Example 2.

[0032] The resulting magnetite concentrate contained 60.9% TFe, 0.20% Sn, and an iron recovery of 80.7%. The non-magnetic mineral Sn was 0.71%. Under these conditions, the cassiterite activation reaction was incomplete, and the calcium stannate phase was not effectively formed. Tin was mostly present as unreacted SnO2 or in an amorphous state, with small grains encased in the gangue. This resulted in poor responsiveness in subsequent gravity separation and flotation, a significant decrease in separation performance, and ineffective enrichment.

[0033] Comparative Example 3

[0034] The only difference between this comparative example and Example 3 is that the concentration of the calcination atmosphere [CO / (CO+CO2)] is higher and the volume fraction of C is 30%. For other operations and parameters, refer to Example 3.

[0035] The resulting magnetite concentrate had a TFe content of 57.9%, a Sn content of 0.71%, and an Fe recovery of 58.5%. Sn enrichment in the non-magnetic product was not significant, and the tin recovery rate decreased. This suggests that an overly strong CO atmosphere can lead to excessive iron reduction, hindering subsequent gravity separation and flotation. Tin failed to form a stable complex, resulting in poor flotation selectivity.

[0036] Comparative Example 4

[0037] The only difference between this comparative example and Example 4 is that the reselection step is eliminated, and other operations and parameters are the same as those in Example 4.

[0038] The Sn content of the flotation concentrate was 1.3%, that of the tailings was 0.6%, and the tin recovery rate was 41.2%. This indicates that the gangue interference was large without gravity separation, the flotation effect was significantly reduced, and the tin recovery rate and grade were both poor.

[0039] Comparative Example 5

[0040] The only difference between this comparative example and Example 5 is that the flotation step is omitted, and other operations and parameters are the same as those in Example 5.

[0041] The Sn content of the heavy product obtained was 1.8%, accompanied by a significant amount of calcified impurities, and the tin recovery rate was 46.3%. This indicates that gravity separation alone is difficult to obtain a high-tin concentrate and requires flotation to further improve recovery efficiency.

[0042] Comparative Example 6

[0043] The only difference between this comparative example and Example 1 is that only fatty acid collectors are used, and no hydroxamic acid collector is compounded. Other operations and parameters are the same as those in Example 1.

[0044] Flotation results showed that the Sn grade of the tin-rich concentrate was 2.9%, the Sn content of the tailings was 0.19%, and the tin recovery rate was only 64.7%. This indicates that without the addition of hydroxamic acid, the fatty acid capture capacity was limited, the selectivity was poor, and some calcium stannate was not effectively enriched.

[0045] Example 1

[0046] Tin-iron tailings (TFe 44.8%, Sn 0.46%) and calcium carbonate (CaCO₃ ≥98%) from a mining area were mixed in a mass ratio of 1:0.3 and ground to a particle size of -0.1 mm. 12% water and 2% bentonite were added to the mixture and granulated to produce 7-10 mm agglomerates. The calcination temperature was set at 900°C for 45 minutes, and the ratio of [CO₂ / (CO₂+CO₂)] × 100% = 15% was used. After cooling, the calcined product was ball-milled to a particle size of -0.074 mm, accounting for 90% of the total particle size. Magnetic separation was performed at a magnetic field strength of 900 GS, yielding a magnetite concentrate with a TFe content of 64.3%, a Sn content of 0.06%, and an Fe recovery of 85.6%. After magnetic separation, the non-magnetic fraction was subjected to centrifugal gravity separation (15 Hz, 8 LPM backwash water) to initially enrich a tin-containing concentrate of 1.8%. The tin-containing concentrate was then subjected to flotation. The pulp concentration was first adjusted to 36%, and the pH was adjusted to 9.5 with sulfuric acid. 600 g / t of a composite collector (fatty acid:hydroxamic acid = 3:1) was then added and stirred for 3 minutes. 20 g / t of a frother, terpineol, was then added and stirred for 2 minutes. The concentrate then underwent a roughing, scavenging, and finishing process (roughing for 3 minutes, scavenging for 2 minutes, and finishing for 2 minutes). A tin-rich concentrate with a Sn content of 4.4% and a tin recovery of 83.4% was obtained. Overall, the tin-iron separation was excellent.

[0047] Example 2

[0048] Tin-iron tailings (TFe 43.5%, Sn 0.51%) from a mineral processing tailings were mixed with calcite (CaCO₃ ≥95%) at a mass ratio of 1:0.3, and then 12% water and 2% bentonite were added to produce 8-10mm agglomerates. The dried agglomerates were calcined in a CO / (CO+CO₂) atmosphere with a volume fraction of 15% at 925°C for 60 minutes. The calcined product was cooled and ball milled, resulting in a -0.074mm particle size fraction of 90%. After 1000GS magnetic separation, a magnetite concentrate with a TFe content of 65.1%, Sn 0.05%, and an Fe recovery of 86.2% was obtained. The non-magnetic fraction was subjected to centrifugal gravity separation (16Hz, backwash water 9LPM), yielding 1.9% tin-bearing minerals for further flotation. The ore pulp was prepared at 38%, the pH adjusted to 9 with sulfuric acid, and a 3:1 mixture of fatty acid and hydroxamic acid was added as a collector (600g / t) and stirred for 3 minutes. A frother (20g / t of terpineol) was added and stirred for 2 minutes. The ore was subjected to a roughing cycle of 3 minutes, a scavenging cycle of 2 minutes, and a finishing cycle of 2 minutes. The resulting tin-rich concentrate contained 4.1% Sn, and the tailings contained 0.08% Sn, for a tin recovery rate of 86.3%.

[0049] Example 3

[0050] A low-sulfur tin-iron tailings (TFe 46.1%, Sn 0.48%) was mixed with limestone (CaCO3 ≥97%) at a mass ratio of 1:0.3. 10% water and 2% bentonite were added to form granules (8-10 mm). The calcination temperature was 875°C for 50 minutes, and the CO atmosphere was 15% by volume. The roasted ore was ball-milled to a particle size of -0.074 mm, accounting for 90%, and then fed into a magnetic separator with a magnetic field strength of 1200 GS to obtain a magnetite concentrate with a TFe of 62.7% and a Sn of 0.07%. The non-magnetic matter was subjected to centrifugal gravity separation at a frequency of 18 Hz and a backwash water of 8 LPM. The gangue minerals were discarded to obtain a rougher ore. The rougher ore was flotated under the conditions of a pulp concentration of 37%, a pulp adjusted to pH = 10 with sulfuric acid, and a mixture of fatty acid and hydroxamic acid (3:1) as a collector. 600 g / t of collector was added and stirred for 3 minutes, and then 20 g / t of frother terpineol was added and stirred for 2 minutes. After a roughing process of 3 minutes, a scavenging process of 2 minutes, and a concentrating process of 2 minutes, a tin-rich concentrate with a Sn of 4.2% and a tin recovery rate of 82.7% was obtained, and the process adaptability was good.

[0051] Example 4

[0052] Using a tin-iron tailings as raw material (TFe 42.7%, Sn 0.43%) as raw material, it was mixed with calcium carbonate (CaCO3 ≥98%) at a ratio of 1:0.25, 11% water and 2.3% bentonite to prepare 8-10mm particle size agglomerates, which were roasted at 900℃ and a CO / (CO+CO2) volume fraction of 12% for 45 minutes. The calcined product was ball-milled to a -0.074mm particle size, accounting for 90% of the total particle size. The product was then subjected to magnetic separation in an 1100GS magnetic separator, yielding a magnetite concentrate with a TFe content of 63.2%, Sn of 0.04%, and an Fe recovery of 82.3%. The non-magnetic material was then subjected to rougher separation at a 17Hz frequency and 7.5LPM backwash water. The rougher separation was performed using a 36% slurry concentration, a pH of 9.5 adjusted with sulfuric acid, and a 3:1 mixture of fatty acid and hydroxamic acid as a collector. 600g / t of the combined collector (fatty acid:hydroxamic acid = 3:1) was added and stirred for 3 minutes. Then, 20g / t of terpineol, a foaming agent, was added and stirred for 2 minutes. The product was subjected to a 3-minute roughing cycle, a 2-minute scavenging cycle, and a 2-minute finishing cycle. The resulting tin-rich concentrate had a Sn grade of 3.6% and a tin recovery of 85.5%.

[0053] Example 5

[0054] Tin-iron tailings (TFe 40.2%, Sn 0.39%) from a certain mine were used as raw materials and mixed with limestone (CaCO3 ≥97%) at a ratio of 1:0.3, with 11% water and 2.3% bentonite to prepare 8-10 mm agglomerates. The dried agglomerates were calcined in a mixed atmosphere with a CO volume fraction of 15% at a temperature of 925°C for 35 min. After ball milling to a particle size of -0.074mm, accounting for 90%, the grinding products were magnetically separated in a magnetic separator with a magnetic field strength of 900GS to obtain a magnetite concentrate with a TFe content of 60.7% and a Sn content of 0.05%. The non-magnetic material was subjected to rough separation under the conditions of frequency 16HZ and backwash water 8LPM. The rougher ore was adjusted to pH=9.5 with sulfuric acid, and a mixture of fatty acid and hydroxamic acid (3:1) was used as a collector. Flotation was carried out at a pulp concentration of 38%. The flotation process was as follows: first, 600g / t of collector was added and stirred for 3 minutes, then 20g / t of frother terpineol was added and stirred for 2 minutes. After a roughing process of 3 minutes, a scavenging process of 2 minutes, and a concentrating process of 2 minutes, a tin-rich concentrate with a Sn content of 3.2% was obtained, and the tin recovery rate was 78.9%.

Claims

1. A method for preparing magnetite concentrate and tin-rich concentrate from tin-iron tailings, characterized in that: The tin-iron tailings and calcium carbonate minerals are mixed and agglomerated, and the resulting agglomerates are placed in a mixed atmosphere of CO and CO2 for reduction roasting to obtain a roasted product containing calcium stannate and magnetite; the roasted product is ball-milled and slurried, and then subjected to magnetic separation to recover magnetite concentrate, gravity separation to remove gangue, and flotation to recover calcium stannate concentrate.

2. The method for preparing magnetite concentrate and tin-rich concentrate from tin-iron tailings according to claim 1, characterized in that: The mass ratio of the tin-iron tailings to the calcium carbonate minerals is 1:0.2-0.

4.

3. A method for preparing magnetite concentrate and tin-rich concentrate from tin-iron tailings according to claim 1 or 2, characterized in that: The mass content of TFe in the tin-iron tailings is 30% to 55%, and the mass content of Sn is 0.2% to 0.8%; The calcium carbonate mineral includes at least one of industrial-grade calcium carbonate, limestone, and calcite, and its CaCO3 mass content is above 95%.

4. The method for preparing magnetite concentrate and tin-rich concentrate from tin-iron tailings according to claim 1, characterized in that: The reduction roasting conditions are as follows: the atmosphere is a mixed atmosphere of CO and CO2 with a CO volume concentration of 5% to 17%, the temperature is 850° C. to 950° C., and the time is 30 to 60 minutes.

5. The method for preparing magnetite concentrate and tin-rich concentrate from tin-iron tailings according to claim 1, characterized in that: During the ball milling process, the mass proportion of the -0.074 mm particle size is controlled to be more than 90%.

6. The method for preparing magnetite concentrate and tin-rich concentrate from tin-iron tailings according to claim 1, characterized in that: The magnetic field strength used in the magnetic separation process is 800-1200 Gs.

7. The method for preparing magnetite concentrate and tin-rich concentrate from tin-iron tailings according to claim 1, characterized in that: The high-frequency centrifugal gravity separation method is adopted in the gravity separation process. The frequency of the high-frequency centrifugal gravity separation is 15-20 Hz, and the flow rate of the backwash water is controlled to be 7-10 LPM.

8. The method for preparing magnetite concentrate and tin-rich concentrate from tin-iron tailings according to claim 1, characterized in that: During the flotation process, the pulp mass concentration is adjusted to 35-40%, and the pH is adjusted to 9-10.

9. A method for preparing magnetite concentrate and tin-rich concentrate from tin-iron tailings according to claim 1 or 8, characterized in that: During the flotation process, a calcium stannate mineral collector is used, which is composed of a fatty acid collector and a hydroxamic acid collector in a mass ratio of 2 to 4:1.

Images