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

Abstract

The present invention discloses a method for preparing magnetite concentrate and tin-rich concentrate from tin-iron tailings, belonging to the technical field of tailings resource utilization. The method comprises 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; the roasted product is subjected to ball milling and slurry mixing, followed by magnetic separation to recover magnetite concentrate, gravity separation to remove gangue, and flotation to recover calcium stannate concentrate. This method not only achieves efficient conversion of tin-iron tailings at a relatively low temperature, but also achieves efficient enrichment and separation of tin and iron minerals, and obtains high-grade calcium stannate concentrate and magnetite products, truly realizing efficient, clean, and large-scale utilization of tin-iron tailings resources.

Description

Technical Field

[0001] This invention relates to a method for processing tin-iron tailings, and particularly to a method for preparing magnetite concentrate and tin-rich concentrate from tin-iron tailings, belonging to the field of comprehensive utilization technology of non-ferrous metal resources. Background Art

[0002] Tin-iron tailings are a type of complex and difficult-to-process ore with relatively abundant reserves but low levels of development and utilization. These ores typically contain 30%–55% iron and 0.1%–0.5% tin, with some high-grade ores containing over 0.8% tin. Because they are rich in iron and also contain economically valuable tin, they possess strong resource utilization potential. However, due to their complex mineral composition and difficulty in processing, they have not been effectively developed for a long time.

[0003] Tin in tin-iron tailings mainly exists in the form of cassiterite (SnO2), exhibiting an extremely fine-grained dispersive state. It often exists as individual particles or closely associated with iron oxides (such as hematite and limonite), resulting in a dense structure that is difficult to effectively liberate using conventional physical beneficiation methods. This complex mineral dispersive relationship leads to a blurred interface between tin and iron, resulting in "co-floating" or "co-sinking" phenomena during beneficiation. This results in poor selectivity for tin-iron separation, limited separation indicators, and seriously affects the comprehensive recovery of resources.

[0004] For the treatment of tin-iron tailings, pyrometallurgical methods are currently the main means of separating tin and iron. Typical methods include sulfidation volatilization, chlorination roasting, and weak reduction volatilization. Among them, sulfidation volatilization effectively removes tin by generating SnS, which volatilizes at high temperatures, but the operating temperature usually needs to reach 1180℃~1300℃, accompanied by a large amount of SO2 gas emission, which places a heavy burden on equipment and the environment. Chlorination roasting relies on the reaction of chlorinating agents with tin and iron to generate volatile chlorides, achieving tin volatilization and recovery at around 1000℃, but it suffers from severe equipment corrosion and serious chlorine pollution. Weak reduction volatilization utilizes the high vapor pressure of SnO at high temperatures to achieve volatilization, but it has stringent requirements for atmosphere and temperature control and is not conducive to iron recovery.

[0005] Currently, patented technologies utilize low-temperature reducing roasting combined with magnetic separation and leaching for the comprehensive utilization of tin-iron ore. For example, Chinese patent (CN104152671B) proposes using sodium carbonate as an additive to roast tin-iron ore under a weak 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 can be obtained separately. This method has a low roasting temperature (850℃~945℃), low energy consumption, good tin-iron separation effect, and is a green and environmentally friendly process. However, the recovery of residual tin in the tailings requires a wet process, which involves many steps, is cumbersome, and the tin-rich product has an unstable form, posing certain difficulties for subsequent resource utilization. Summary of the Invention

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

[0007] To achieve the above-mentioned technical objectives, the present invention provides a method for preparing magnetite concentrate and tin-rich concentrate from tin-iron tailings. The method involves mixing and agglomerating tin-iron tailings with calcium carbonate minerals, and then subjecting the resulting agglomerated granules to reduction roasting under a mixed atmosphere of CO and CO2 to obtain a roasted product containing calcium stannate and magnetite. The roasted product is then ball-milled and slurry-adjusted, and subsequently subjected to magnetic separation to recover magnetite concentrate, gravity separation to remove gangue, and flotation to recover calcium stannate concentrate.

[0008] The key to this invention lies in: firstly, using calcium carbonate as an additive to achieve low-temperature conversion of tin-iron tailings, transforming them into calcium stannate and magnetite minerals; then, using magnetic separation, gravity separation, and flotation processes to separate the calcium stannate and magnetite minerals, thereby obtaining high-grade calcium stannate concentrate and magnetite concentrate. More specifically, this invention uses calcium carbonate as an additive, and the tin-iron tailings are subjected to reduction roasting in a mixed atmosphere of CO and CO2. The reduction roasting process converts the high-valence iron oxides in the tin-iron tailings into magnetite and activates cassiterite. Simultaneously, calcium carbonate decomposes at high temperature to form active calcium oxide. Active calcium oxide plays a crucial role in the roasting process; on the one hand, it acts as a flux to lower the roasting temperature of the tin-iron tailings; on the other hand, it acts as a tin oxide converter, capturing and converting cassiterite into calcium stannate minerals. The transformation of tin and iron into magnetite and calcium stannate minerals through reduction roasting provides favorable conditions for subsequent mineral separation processes. In the mineral separation process, the strong magnetic properties of magnetite are preferentially utilized to first separate it from calcium stannate minerals through magnetic separation, which can obtain high-grade magnetite concentrate. However, the grade of calcium stannate minerals in the magnetic separation tailings is relatively low. This invention adopts a method of gravity separation tailings removal combined with flotation enrichment to recover calcium stannate minerals. After centrifugal gravity separation, most gangue minerals can be efficiently removed, achieving preliminary enrichment of calcium stannate minerals and improving the feed grade and separation boundary of subsequent flotation. On this basis, further tin-enriched concentrate of calcium stannate minerals is obtained through flotation, resulting in high-grade calcium stannate minerals.

[0009] As a preferred embodiment, the mass ratio of the tin-iron tailings to the calcium carbonate minerals is 1:0.2-0.4. The ratio of tin-iron tailings to calcium carbonate minerals is best controlled within the preferred range. If the proportion of calcium carbonate is too low, it will result in a low cassiterite conversion rate and incomplete recovery, while also requiring a high roasting temperature. Conversely, if the proportion of calcium carbonate is too high, some iron minerals will be converted into minerals such as calcium ferrite that cannot be recovered through magnetic separation, resulting in a low iron recovery rate.

[0010] As a preferred embodiment, the tin-iron tailings contain 30%–55% TFe and 0.2%–0.8% Sn by mass. Tin-iron tailings are characterized by high iron content and low tin content, with tin mainly existing in the form of cassiterite (SnO2) in an extremely fine-grained disseminated state. They often exist as single particles or closely coexist with iron oxides (such as hematite and limonite), resulting in a dense structure.

[0011] As a preferred embodiment, the calcium carbonate mineral includes at least one of industrial-grade calcium carbonate, limestone, and calcite, with a CaCO3 content of 95% or higher. The calcium carbonate mineral can be derived from industrial by-product calcium carbonate, mineral processing tailings, or industrial analytical grade reagents. The addition of calcium carbonate minerals serves two purposes: firstly, to adjust the mineral alkalinity, and secondly, to enhance the activation reaction behavior of cassiterite.

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

[0013] As a preferred embodiment, the reduction roasting conditions are: a mixed atmosphere of CO and CO2 with a CO volume concentration of 5%–17%, a temperature of 850℃–950℃, and a time of 30–60 minutes. Under the synergistic control of the preferred roasting temperature and atmosphere conditions, the reduction of high-valence iron minerals can be achieved. For example, hematite, limonite, and other minerals can be reduced and transformed into magnetite minerals under a weak reducing atmosphere. Simultaneously, the decomposition of calcium carbonate minerals and the conversion of active calcium oxide and cassiterite into calcium stannate minerals can be achieved. If the roasting temperature is too low, the tin-iron tailings will not be completely converted; if the roasting temperature is too high, the iron minerals will be converted into new phases such as calcium ferrite, reducing the iron recovery rate. Furthermore, the reducing atmosphere affects the magnetization of high-valence iron minerals and the activation of cassiterite. Too low a CO volume concentration will result in incomplete magnetization of high-valence iron minerals and insufficient activation of cassiterite, reducing its recovery rate. Too high a CO volume concentration will easily cause some cassiterite to be reduced, which is detrimental to tin recovery. By reducing roasting, cassiterite can be converted into calcium stannate minerals that are easy to float, while hematite and limonite are reduced to magnetite, thus achieving a difference in the mineral phase properties of the two.

[0014] As a preferred embodiment, the ball milling process aims to maintain a mass percentage of -0.074mm particles of 90% or higher. Grinding to an appropriate particle size facilitates subsequent mineral separation processes such as magnetic separation, gravity separation, and flotation.

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

[0016] As a preferred embodiment, the gravity separation process employs high-frequency centrifugal gravity separation at a frequency of 15-20 Hz, with the backwash water flow rate controlled at 7-10 LPM. The non-magnetic minerals after magnetic separation contain calcium stannate minerals, primarily calcium stannate, as well as small amounts of unreacted cassiterite and gangue. High-frequency centrifugal gravity separation is used for preliminary density classification. Under these preferred gravity separation conditions, efficient removal of gangue minerals and enrichment of tin-containing minerals can be achieved, improving the subsequent flotation feed grade and separation boundary, and increasing the tin enrichment factor. Under these conditions, gangue inclusions such as quartz and calcite with a particle size range of 0.045-0.25 mm can be stably removed, and the tin grade in the gravity-separated ore can be increased to over 1.7%.

[0017] As a preferred embodiment, during the flotation process, the pulp concentration is adjusted to 35-40%, and the pH is adjusted to 9-10. Under the influence of the preferred pulp concentration and pH, it is beneficial to maintain the stability of the surface charge properties of calcium stannate and the adsorption and activation of the collector.

[0018] As a preferred embodiment, the flotation process employs a calcium stannate mineral collector composed of a fatty acid collector and a hydroxamic acid collector at a mass ratio of 2-4:1. The fatty acid collector and the hydroxamic acid collector are used in appropriate proportions. The fatty acid (such as oleic acid or its sodium salt) provides basic hydrophobic adsorption capacity, while the hydroxamic acid (such as sodium phenylhydroxamic acid) enhances the selective complexation capacity for tin-containing phases such as calcium stannate. The synergistic effect of these two components significantly improves tin recovery efficiency and flotation concentrate quality. Ultimately, high-grade, low-tin magnetite concentrate and high-tin-grade tin-rich concentrate are obtained, significantly improving the overall resource recovery rate and optimizing economic benefits and environmental indicators. In the flotation system, the fatty acid collector establishes initial hydrophobicity, while the hydroxamic acid brittle molecules enhance collecting power and selectivity through chelation, effectively suppressing the co-flotation of associated iron and calcium impurities, achieving high selective enrichment of tin, and ultimately obtaining a tin-rich concentrate with a tin grade of 3%-5%. The flotation process includes one roughing, one cleaning, and one sweeping.

[0019] In the process of agglomerating tin-iron tailings and calcium carbonate minerals according to the present invention, 8%–15% water and 1%–3% binder (such as bentonite or hydrated lime) are added for granulation or briquetting. This produces 8–10 mm briquettes, which is beneficial for gas heat transfer and uniform 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 sulfidation volatilization method can achieve high-temperature volatilization and recovery of tin, most of the iron elements are lost with the slag, making it impossible to obtain high-grade iron concentrate; while conventional flotation or gravity separation methods do not completely separate tin and iron, resulting in high tin content in magnetite, which is difficult to meet smelting requirements. This invention achieves effective decoupling of cassiterite and iron oxide in structure through controlled reduction roasting under a CO / CO2 atmosphere. The tin and iron content of the magnetite obtained by magnetic separation after roasting fully meets the blast furnace raw material standards, while 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) This invention takes a specific process flow as the main line. Compared with complex processes such as multi-stage wet leaching and multi-stage extraction to extract tin, the process links are clear, the parameter control range is wide (such as CO concentration and roasting temperature), it does not rely on highly corrosive chemicals or complex reagent systems, has high industrial operability, and is adapted to the characteristics of strong fluctuations in different tin and iron tailings raw materials.

[0023] (3) Compared with traditional processing methods, the present invention can not only effectively extract high-grade magnetite, but also achieve deep enrichment of non-magnetic tin resources, and minimize the loss rate of tin in tailings. Attached Figure Description

[0024] Figure 1 This is a process flow diagram of the present invention for preparing magnetite concentrate and tin-rich concentrate from tin-iron tailings. Detailed Implementation

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

[0026] In the following specific embodiments and comparative embodiments, the proportion of tin-iron tailings and calcium carbonate ground to -0.1mm particle size is 100%.

[0027] Comparative Example 1

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

[0029] The obtained magnetite concentrate had a TFe content of 58.6%, a Sn content of 0.26%, and an iron recovery rate of 78.9%. Sn was not significantly enriched in the non-magnetic materials, with a Sn content of 0.59%. The results indicate that the absence of calcium carbonate resulted in the failure of cassiterite activation and transformation; tin was mainly contained within the magnetic products, and tin-iron separation was inadequate.

[0030] Comparative Example 2

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

[0032] The obtained magnetite concentrate had a TFe content of 60.9%, a Sn content of 0.20%, and an iron recovery rate of 80.7%; the non-magnetic mineral Sn content was 0.71%. Under these conditions, the activation reaction of cassiterite was incomplete, the calcium stannate phase was not effectively formed, and tin was mostly present as unreacted SnO2 or in an amorphous state. The grains were small and encapsulated in the gangue, resulting in poor response to subsequent gravity separation and flotation, a significant decrease in separation indicators, and ineffective enrichment.

[0033] Comparative Example 3

[0034] The only difference between this comparative example and Example 3 is that the concentration of the roasting atmosphere [CO / (CO+CO2)] is higher and the volume fraction of C is 30%. Other operations and parameters are the same as in Example 3.

[0035] The obtained magnetite concentrate had a TFe content of 57.9%, a Sn content of 0.71%, and an Fe recovery rate of 58.5%. Sn enrichment was not significant in the non-magnetic products, and the tin recovery rate decreased. This indicates that an excessively strong CO atmosphere easily leads to over-reduction of the iron phase, which is unfavorable for 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 canceled; other operations and parameters are the same as in Example 4.

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

[0039] Comparative Example 5

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

[0041] The obtained heavy product Sn was 1.8%, accompanied by a large amount of calcified impurities; the tin recovery rate was 46.3%. This indicates that gravity separation alone is insufficient to obtain high tin enrichment, and combined flotation is needed 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, without the addition of hydroxamic acid collectors. Other operations and parameters are the same as in Example 1.

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

[0045] Example 1

[0046] Using tin-iron tailings (TFe 44.8%, Sn 0.46%) and calcium carbonate (CaCO3 ≥98%) from a certain mining area as raw materials, they were mixed at a mass ratio of 1:0.3 and ground separately to a particle size of -0.1 mm. 12% water and 2% bentonite were added for granulation, yielding 7-10 mm agglomerates. The roasting temperature was set at 900℃, and the roasting time was 45 min, with [CO / (CO+CO2)]×100%=15%. After cooling, the roasted product was ball-milled to a particle size of -0.074 mm, with 90% of the product reaching this size. The magnetic separation field strength was 900 GS, yielding a magnetite concentrate with 64.3% TFe, 0.06% Sn, and an Fe recovery rate of 85.6%. The non-magnetic fraction after magnetic separation was subjected to centrifugal gravity separation (15 Hz, 8 LPM backwash water) to initially enrich a 1.8% tin-containing concentrate. This tin-containing concentrate underwent flotation. The pulp concentration was first adjusted to 36%, and the pH was adjusted to 9.5 using sulfuric acid. Then, 600 g / t of a compound collector (fatty acid: hydroxamic acid = 3:1) was added and stirred for 3 minutes. Next, 20 g / t of frother, terpineol, was added and stirred for 2 minutes. The mixture then underwent a roughing, scavenging, and cleaning process: roughing for 3 minutes, scavenging for 2 minutes, and cleaning for 2 minutes. This yielded a tin-rich concentrate with 4.4% Sn and a tin recovery rate of 83.4%. Overall, the tin-iron separation effect was good.

[0047] Example 2

[0048] Using tin-iron tailings (TFe 43.5%, Sn 0.51%) from a mineral processing plant as raw material, calcite (CaCO3 ≥95%) was mixed at a mass ratio of 1:0.3, and 12% water and 2% bentonite were added to form 8-10mm particle size agglomerates. The dried agglomerates were calcined in a mixed atmosphere with a CO / (CO+CO2) volume fraction of 15% at 925℃ for 60 minutes. After cooling, the calcined product was ball-milled. After ball milling, the -0.074mm particle size accounted for 90%. Magnetite concentrate was obtained by 1000GS magnetic separation with a TFe content of 65.1%, Sn content of 0.05%, and Fe recovery rate of 86.2%. The non-magnetic fraction was centrifuged and gravity separated (16Hz, 9 LPM backwash water) to obtain 1.9% tin-bearing minerals for further flotation. The pulp was prepared with a 38% ore content, and the pH was adjusted to 9 with sulfuric acid. A 3:1 mixture of fatty acids and hydroxamic acid collector (600 g / t) was added and stirred for 3 min. Then, 20 g / t of frother and a suitable amount of terpineol were added and stirred for 2 min. The mixture underwent a three-stage roughing process (3 min), a two-stage scavenging process (2 min), and a two-stage cleaning process (2 min). The resulting tin-rich concentrate had a Sn content of 4.1%, the tailings had a Sn content of 0.08%, and the tin recovery rate reached 86.3%.

[0049] Example 3

[0050] Using a low-sulfur tin-iron tailings ore (TFe 46.1%, Sn 0.48%) as raw material, it was mixed with limestone (CaCO3 ≥97%) at a mass ratio of 1:0.3, and 10% water and 2% bentonite were added for granulation (8-10 mm). The roasting temperature was 875℃, the roasting time was 50 min, and the CO atmosphere volume fraction was 15%. Roasted minerals were ball-milled to a particle size of -0.074mm (90%) and fed into a magnetic separator with a magnetic field strength of 1200GS to obtain a magnetite concentrate with TFe of 62.7% and Sn of 0.07%. Non-magnetic materials were centrifuged and gravity separated at a frequency of 18Hz with backwash water of 8LPM to remove gangue minerals and obtain rougher ore. The rougher ore was subjected to flotation under the conditions of a pulp concentration of 37%, a pulp pH adjusted to 10 with sulfuric acid, and a mixture of fatty acids and hydroxamic acid (3:1) as a collector. 600g / t of collector was added and stirred for 3min, followed by the addition of 20g / t of frother terpineol and stirring for 2min. After a roughing process of 3min, a scavenging process of 2min, and a cleaning process of 2min, a tin-rich concentrate with Sn of 4.2% and a tin recovery rate of 82.7% was obtained, demonstrating good process adaptability.

[0051] Example 4

[0052] Using a certain tin-iron tailings as raw material (TFe 42.7%, Sn 0.43%), it was mixed with calcium carbonate (CaCO3 ≥98%) at a ratio of 1:0.25, with 11% water and 2.3% bentonite to form 8-10mm particle size agglomerates. The agglomerates were then roasted for 45 minutes at 900℃ and a CO / (CO+CO2) volume fraction of 12%. The roasted product was ball-milled to a particle size of -0.074 mm (90% of the total) and then subjected to magnetic separation in a 1100 GS magnetic separator to obtain magnetite concentrate with 63.2% TFe, 0.04% Sn, and an Fe recovery rate of 82.3%. The non-magnetic material was subjected to gravity separation at a frequency of 17 Hz and a backwash water flow rate of 7.5 LPM to obtain rougher ore. The rougher ore was prepared with a pulp concentration of 36%, a pulp pH adjusted to 9.5 with sulfuric acid, and a 3:1 mixture of fatty acids and hydroxamic acid as the collector. First, 600 g / t of the mixed collector (fatty acid:hydroxamic acid = 3:1) was added and stirred for 3 min. Then, 20 g / t of frother terpineol was added and stirred for 2 min. The mixture underwent a 3-min roughing, a 2-min scavenging, and a 2-min cleaning process to obtain tin-rich concentrate with a Sn grade of 3.6% and a tin recovery rate of 85.5%.

[0053] Example 5

[0054] Using tin-iron tailings (TFe 40.2%, Sn 0.39%) from a certain mine as raw material, limestone (CaCO3 ≥97%) was mixed with 11% water and 2.3% bentonite at a ratio of 1:0.3 to prepare 8-10 mm agglomerates. The dried agglomerates were calcined in a mixed atmosphere with a CO volume fraction of 15% at a calcination temperature of 925℃ for 35 min. The ore was ball-milled to a particle size of -0.074 mm (90% of the total particle size). The grinding product was then magnetically separated in a magnetic separator with a magnetic field strength of 900 GS 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 gravity separation at a frequency of 16 Hz and a backwash water flow rate of 8 LPM to obtain a rougher. The rougher was then subjected to flotation with a pulp adjusted to pH 9.5 with sulfuric acid, using a mixture of fatty acids and hydroxamic acid (3:1) as the collector, and a pulp concentration of 38%. The flotation process involved adding 600 g / t of collector and stirring for 3 min, followed by adding 20 g / t of frother terpineol and stirring for 2 min. After a 3 min roughing stage, a 2 min scavenging stage, and a 2 min cleaning stage, a tin-rich concentrate with a Sn content of 3.2% and a tin recovery rate of 78.9% was obtained.

Claims

1. A method for preparing magnetite concentrate and tin-rich concentrate from tin-iron tailings, characterized in that: Tin-iron tailings and calcium carbonate minerals are mixed and agglomerated. The resulting agglomerates are then subjected to reduction roasting in a mixed atmosphere of CO and CO2 to obtain a roasted product containing calcium stannate and magnetite. The roasted product is then ball-milled and slurry-adjusted, and then subjected to magnetic separation to recover magnetite concentrate, gravity separation to remove gangue, and flotation to recover calcium stannate concentrate. The reselection process employs a high-frequency centrifugal reselection method, with a frequency of 15~20Hz and a backwash water flow rate of 7~10LPM. In 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.

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 to 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 tin-iron tailings contained 30%–55% Tfe and 0.2%–0.8% Sn by mass. The calcium carbonate minerals include at least one of industrial-grade calcium carbonate, limestone, and calcite, with a CaCO3 content of 95% or more.

4. The method for preparing magnetite concentrate and tin-rich concentrate from tin-iron tailings according to claim 1, characterized in that: The conditions for the reduction roasting are as follows: the atmosphere is a mixture of CO and CO2 with a CO volume concentration of 5% to 17%, the temperature is 850℃ to 950℃, 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 percentage of particles with a particle size of -0.074 mm is controlled to be above 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. A 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 concentration is adjusted to 35-40%, and the pH is adjusted to 9-10.

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