Mining control method based on coal mine digging, mining, filling and retaining coupling

Abstract

The invention discloses a mining control method based on coal mine excavation and mining filling and retaining coupling, and belongs to the technical field of coal mine mining control. The method comprises the following steps: constructing a physical similar model for simulating geological conditions of a target coal mine, excavating a simulated roadway according to a preset tunneling sequence to form a simulated mining surface, implementing simulated filling operation to form a simulated filling body, and reserving simulated coal pillars between the simulated roadway and the simulated mining surface to form a supporting structure; in the mining and filling process, overlying strata and ground surface subsidence displacement data are collected, a roof subsidence curve is drawn, wave crest and wave trough positions are identified, overlying strata movement differences are analyzed, the ground surface subsidence rate is calculated, and on-site actual mining parameters are guided accordingly; according to the method, the supporting effect can be optimized, the overlying strata movement rule can be accurately mastered, and the mining safety problem caused by insufficient data analysis and unreasonable supporting design in coal mine excavation and mining is solved; scientific guidance is provided for on-site mining parameter adjustment, and coal mining requirements under complex geological conditions are met.

Description

A mining control method based on the coupling of coal mine excavation, mining, filling and retention. Technical Field

[0001] This invention belongs to the field of coal mine mining control technology, specifically a mining control method based on the coupling of coal mine excavation, mining, filling and retention. Background Technology

[0002] In coal mining, controlling overburden and surface subsidence, and ensuring mining safety are core requirements. Current technologies generally employ the construction of physically similar models to simulate mining scenarios, using simulated tunneling and backfilling operations, combined with displacement monitoring data to guide on-site mining. In conventional backfilling mining techniques, support methods are mostly single backfill bodies or single coal pillars. While some technologies utilize both backfill bodies and coal pillars, the spatial relationship between the two is not rationally planned, failing to create a synergistic support effect. In the displacement monitoring and data analysis stage, current technologies only collect and simply record overburden and surface subsidence displacement data, without in-depth data mining, making it difficult to accurately grasp the inherent patterns of overburden movement.

[0003] The unreasonable design of the support structure leads to insufficient overburden support strength, significant overburden movement, and prominent surface subsidence, which in turn affects the surrounding environment and mining safety. Simultaneously, the low utilization rate of displacement data makes it impossible to clearly define the differences in overburden movement at corresponding locations of simulated coal pillars and simulated mining faces through data analysis. It also makes it difficult to establish a correlation between surface subsidence and simulated mining thickness, resulting in a lack of precise guidance for adjusting on-site mining parameters and an inability to adapt to mining needs under different geological conditions. Optimizing the support structure to improve overburden control and providing scientific guidance for on-site mining parameters through precise data analysis are urgent problems to be solved in the field of coal mine mining control. Summary of the Invention

[0004] This invention overcomes the shortcomings of existing technologies and proposes a mining control method based on the coupling of coal mine excavation and filling; it solves the mining safety problems caused by insufficient data analysis and unreasonable support design in coal mine excavation.

[0005] This invention is achieved through the following technical solution: a mining control method based on the coupling of coal mine tunneling, mining, filling, and retention, the method comprising: constructing a physical similarity model for simulating the geological conditions of a target coal mine; excavating simulated roadways in the physical similarity model according to a preset tunneling sequence to form a simulated mining face for simulating longwall mining operations, and recording the initial width and spatial position of the simulated mining face; performing simulated filling operations within or beside the simulated mining face according to a preset filling pattern to form a simulated filling body; and retaining a simulated coal pillar between the simulated mining face and the simulated filling body, so that the simulated filling body and the simulated coal pillar are spatially adjacent. To form a supporting structure; during the mining and backfilling operations of the physical similarity model, multiple sets of displacement monitoring points are arranged to collect real-time subsidence displacement data of each overburden stratum and the surface in the physical similarity model; based on the collected subsidence displacement data, the subsidence curve of the roof during mining is plotted, and the morphological characteristics of the subsidence curve are analyzed to identify the peak and trough positions; based on the distribution pattern of the peak and trough positions, the difference in movement of the overburden at the corresponding positions of the simulated coal pillar and the simulated mining face is analyzed, and the ratio of surface subsidence to simulated mining thickness is calculated to obtain the surface subsidence rate; based on the surface subsidence rate, the actual mining parameters on site are guided.

[0006] As a further aspect of the present invention, in the physical similarity model, simulated roadways are excavated according to a preset tunneling sequence to form a simulated mining face for simulating mining operations. This includes: the physical similarity model is scaled down based on the actual borehole columnar section, rock strata mechanical parameters, and mining depth data of the target coal mine; an initial simulated mining area is delineated in the physical similarity model according to the actual mining area design of the target coal mine; a micro excavation tool is used to excavate the first simulated roadway along the coal seam strike within the initial simulated mining area to form the initial simulated mining face; within the initial simulated mining area, subsequent simulated roadways are excavated sequentially, parallel to the first simulated roadway and at preset intervals; the isolation coal walls between adjacent simulated roadways are removed, connecting the independent simulated roadways to form a continuous space with a uniform width, which is the final simulated mining face; throughout the excavation process, the excavation timing, spatial coordinates, and total width of the final simulated mining face of each simulated roadway are recorded.

[0007] As a further aspect of the present invention, the filling mode includes a single-sided filling mode and a double-sided filling mode; in the single-sided filling mode, the simulated filling body is arranged on one side of the simulated mining face; in the double-sided filling mode, the simulated filling body is arranged on both sides of the simulated mining face; a similar material is configured to simulate the actual filling material, the mechanical strength of the similar material being proportionally similar to the actual filling material planned to be used in the target coal mine; when the single-sided filling mode is adopted, at the boundary of the simulated mining face along one dip side, a strip filling of the same length and a set width as the simulated mining face is constructed using the similar material. The simulated filling body is formed by constructing a simulated filling body under the single-sided filling mode. When the double-sided filling mode is adopted, strip-shaped filling bodies of the same length and a set width as the simulated mining face are constructed at the two sides of the dip of the simulated mining face using the similar material, forming the simulated filling body under the double-sided filling mode. During the construction of the simulated filling body, the laying thickness and compaction degree of the similar material are controlled to simulate the roof connection effect and density of the actual filling operation. After the similar material is solidified and stabilized, the simulated filling body is formed, and the specific position, width and relative relationship of the simulated filling body to the simulated mining face are recorded.

[0008] As a further aspect of the present invention, during the excavation of the simulated mining face and the construction of the simulated backfill, a coal seam area that will not be excavated is clearly defined; the coal seam area that will not be excavated is located between the mineable area of ​​the simulated backfill and the simulated mining face, and the coal seam area is completely preserved in the model to form the simulated coal pillar; the width of the simulated coal pillar is measured and recorded, and the width of the simulated coal pillar is the distance from the inner edge of the simulated backfill to the boundary of the simulated mining face along the dip direction of the coal seam; it is ensured that the coal seam of the simulated coal pillar maintains its original connection with the roof and floor strata to simulate the support conditions of the actual coal pillar.

[0009] As a further aspect of the present invention, the step of drawing a subsidence curve of the roof during mining based on the collected subsidence displacement data and analyzing the morphological characteristics of the subsidence curve includes: establishing a two-dimensional coordinate system with the strike direction of the simulated mining face as the abscissa and the subsidence displacement of the overburden as the ordinate; connecting the subsidence displacement data of multiple displacement monitoring points located in the same rock stratum and arranged along the strike direction in the order of the monitoring point positions to draw a roof subsidence curve of the rock stratum; observing the overall shape of the roof subsidence curve, identifying points that are significantly convex relative to the two sides as peaks, and identifying points that are significantly concave relative to the two sides as troughs; comparing the identified peak positions with the spatial positions of the simulated coal pillars in the physical similarity model, and comparing the identified trough positions with the spatial positions of the simulated mining face in the physical similarity model; calculating the difference between the subsidence displacement value at the peak and the subsidence displacement value at the trough, the difference being used to characterize the unevenness of roof subsidence between the coal pillar-supported and mining face areas.

[0010] As a further aspect of the present invention, based on the distribution patterns of the peak and trough positions, the differences in the movement of overburden at the corresponding positions of the simulated coal pillar and the simulated mining face are analyzed, and the surface subsidence rate is calculated. This includes: extracting the final stable subsidence displacement of the displacement monitoring points arranged on the surface layer, calculating the average value of the subsidence displacement of all surface monitoring points, and obtaining the maximum surface subsidence value; obtaining the simulated coal seam thickness value of the simulated mining operation in the physical similarity model; dividing the maximum surface subsidence value by the simulated coal seam thickness value to calculate the surface subsidence rate used to characterize the degree of surface movement; comparing the surface subsidence rate of the single-sided model calculated in the single-sided filling mode with the surface subsidence rate of the double-sided model calculated in the double-sided filling mode; and recording the surface subsidence rate of the single-sided model, the surface subsidence rate of the double-sided model, and their corresponding simulated mining face width and simulated coal pillar width as a related data group.

[0011] As a further aspect of the present invention, the step of guiding actual mining parameters on-site based on the surface subsidence rate includes: after the mining operation of the physical similarity model is completed, the physical similarity model is dissected to observe and measure the damage range of the overburden, and to distinguish the height of the caving zone and the height of the fracture zone; based on the observed height of the caving zone, the height of the fracture zone, and the width of the simulated mining face, the correlation between the width of the mining face and the height of the overburden damage is analyzed; and by combining the analysis results of the surface subsidence rate, the height of the overburden damage, and the width of the mining face, the combination of parameters for guiding the actual mining on-site, including the tunneling width, the arrangement of the backfill body, and the width of the coal pillar, is determined.

[0012] As a further aspect of the present invention, after the mining operation of the physical similarity model is completed, the physical similarity model is cut to observe and measure the damage range of the overburden. This includes: after all displacement monitoring data within the physical similarity model stabilizes, data acquisition is stopped, and preparation is made to perform destructive cutting on the physical similarity model; the physical similarity model is cut along a profile line perpendicular to the simulated mining face to expose the profile of the overburden inside the physical similarity model; on the exposed profile, from the top of the mined coal seam upwards, areas where the rock strata have fractured, collapsed, and lost their layered structure are identified, and the height of the area is the height of the collapse zone; above the collapse zone, areas with obvious delamination and fissures but where the rock strata have not completely collapsed are observed and identified, and the height of the area is the height of the fracture zone; using a measuring tool, the vertical distance from the top of the mined coal seam to the top interface of the collapse zone is directly measured to obtain the measured value of the height of the collapse zone, and the vertical distance from the top of the mined coal seam to the top interface of the fracture zone is also measured to obtain the measured value of the height of the fracture zone.

[0013] As a further aspect of the present invention, the step of analyzing the correlation between the width of the mining face and the overburden failure height based on the observed height of the landslide zone, the height of the fracture zone, and the width of the simulated mining face includes: establishing data pairs between the measured height of the landslide zone obtained under different test schemes and the corresponding width of the simulated mining face, and analyzing the trend of the height of the landslide zone changing with the width of the simulated mining face; establishing data pairs between the measured height of the fracture zone obtained under different test schemes and the corresponding width of the simulated mining face, and analyzing the trend of the height of the fracture zone changing with the width of the simulated mining face; setting a height development threshold for the fracture zone, identifying test schemes where the measured height of the fracture zone is less than the height development threshold, and recording the simulated mining face width conditions corresponding to the test schemes; analyzing the change in the height growth rate of the fracture zone as the simulated mining face width gradually increases, and determining the critical value of the simulated mining face width corresponding to the slowdown in the height growth rate of the fracture zone.

[0014] As a further aspect of the present invention, it also includes the step of optimizing the parameters of the coal pillar and backfill body based on the simulation test results: Based on the calculation results of the surface subsidence rate, a backfill mode that meets the surface movement control target is selected, wherein the backfill mode that meets the surface movement control target is a single-sided backfill mode or a double-sided backfill mode; based on the correlation analysis results between the measured height of the fracture zone and the width of the simulated mining face, the maximum allowable mining face width in actual field mining is determined, wherein the maximum mining face width should control the fracture zone development height below a preset safety threshold; based on the ratio of the measured height of the caving zone to the mining thickness, the preset backfill body strength is verified to effectively support the weight of the rock strata within the caving zone; combined with the experimental phenomenon that the presence of the simulated backfill body can reduce the width of the coal pillar, under the premise of ensuring the stability of the simulated coal pillar and the simulated backfill body, an empirical relationship between the coal pillar width, the backfill body width, and the mining face width is established through fitting multiple sets of experimental data; based on the empirical relationship, the maximum mining face width, and the selected backfill body... The filling mode is used to calculate the recommended actual coal pillar width and backfill layout width. Specifically, this includes: converting the actual geological conditions parameters on site, including actual mining depth, actual mining thickness, and lithological strength, into simulation parameters corresponding to the physical similarity model according to a similarity ratio; setting the mining face width in the converted simulation parameters as the maximum mining face width and substituting it into the empirical formula corresponding to the selected filling mode; using the backfill layout width as an input variable in the empirical formula to solve for the calculated value of the coal pillar width required to achieve stable support; comparing the calculated value of the coal pillar width with the minimum safe coal pillar width calculated based on the on-site rock strength theory, and taking the larger of the two values ​​as the final recommended actual coal pillar width on site; and calculating the matching actual backfill layout width on site based on the final determined actual coal pillar width, the maximum mining face width, and the empirical formula, thus completing the determination of mining control parameters.

[0015] The beneficial effects of this invention compared to existing technologies are as follows: 1. This invention retains a simulated coal pillar between the simulated mining face and the simulated backfill, making the simulated backfill and simulated coal pillar spatially adjacent to each other, forming a collaborative support structure. Compared to the conventional support method of a single backfill or a single coal pillar, the adjacent backfill and coal pillar can achieve synergistic force distribution, disperse the pressure exerted by the overburden, reduce the stress load on a single support structure, and lower the probability of deformation and failure of the support structure. This collaborative support structure can better constrain the movement of the overburden, reduce the deformation amplitude of each layer of the overburden, alleviate surface subsidence, improve the geological environment around the mining area, adapt to mining scenarios under complex geological conditions, and avoid hidden dangers such as roof collapse and roadway deformation caused by insufficient support.

[0016] 2. In the mining and backfilling operations of a physically similar model, this invention deploys multiple sets of displacement monitoring points to collect real-time subsidence and displacement data of each overburden stratum and the surface. It plots the roof subsidence curve during mining, identifies peak and trough positions, analyzes the distribution patterns of peaks and troughs to determine the difference in overburden movement at corresponding positions of the simulated coal pillar and the simulated mining face, and calculates the surface subsidence rate as a percentage of the simulated mining thickness. This surface subsidence rate then guides the actual mining parameters on-site. Compared to conventional techniques that only collect displacement data without in-depth analysis, this method accurately captures the patterns and differences in overburden movement, clarifies the correlation between surface subsidence and simulated mining thickness, and makes adjustments to on-site mining parameters more targeted. It allows for flexible adjustments to parameters such as tunneling sequence and backfilling mode based on the overburden movement and surface subsidence rate in different areas, adapting to the mining needs of different geological conditions, reducing resource waste during mining, and lowering mining safety risks. Attached Figure Description

[0017] Figure 1 is a flowchart of a mining control method based on the coupling of coal mine excavation, mining, filling and retention according to the present invention; Figure 2 is a flowchart of simulated roadway excavation and simulated mining face formation; Figure 3 is a flowchart of simulated coal pillar retention; Figure 4 is a curve comparing the surface subsidence rate under different filling modes; Figure 5 is a curve showing the relationship between the width of the simulated mining face and the height of overburden failure. Detailed Implementation

[0018] To make the technical problems to be solved, the technical solutions, and the beneficial effects of this invention clearer, the invention will be further described in detail with reference to the embodiments and accompanying drawings. It should be understood that the specific embodiments described herein are merely illustrative of the invention and are not intended to limit the invention. The technical solutions of this invention are described in detail below with reference to the embodiments and accompanying drawings, but the scope of protection is not limited thereto.

[0019] Referring to Figure 1, this invention proposes a mining control method based on the coupling of mining, excavation, filling, and retention in coal mines, comprising: constructing a physical similarity model that accurately reflects the geological conditions of the target coal mine. In this physical similarity model, simulated roadways are excavated according to a preset excavation sequence. These roadways are interconnected to form a simulated mining face for simulating actual mining operations. Simultaneously, the initial width of the simulated mining face and its spatial position within the model are recorded. According to a pre-designed filling pattern, simulated filling operations are carried out within or on either side of the simulated mining face to construct a simulated filling body. When forming the simulated mining face and the simulated filling body, a portion of the coal seam that has not been excavated is deliberately reserved between them, i.e., a simulated coal pillar, so that the simulated filling body and the simulated coal pillar are spatially adjacent to each other, jointly forming a support structure. Throughout the entire simulated mining and filling operation, multiple sets of displacement monitoring points are deployed inside the physical similarity model and on the surface to collect real-time subsidence displacement data of each overlying stratum and the surface. Based on the collected subsidence displacement data, a subsidence curve of the roof during mining was plotted. By analyzing the morphological characteristics of this curve, the positions of the peaks and troughs were identified. Based on the distribution patterns of the peaks and troughs, the differences in overburden movement above corresponding positions of the simulated coal pillar and the simulated mining face were analyzed, and the ratio of surface subsidence to the simulated mining thickness, i.e., the surface subsidence rate, was calculated. Based on the obtained surface subsidence rate and other model observation results, key parameters for actual on-site mining were guided and optimized.

[0020] Specifically, in this embodiment, referring to Figure 2, it is assumed that the actual coal seam thickness of the target coal mine is 2.5 meters, the actual mining depth is 300 meters, and the planned process is strip mining followed by backfilling. Based on the actual borehole columnar section, rock strata mechanical parameters, and mining depth data of this target coal mine, the physical similarity model determines a geometric similarity ratio of 1:100 and a bulk density similarity ratio of 1:1.6, from which the strength similarity ratio is calculated. Based on these proportional relationships, materials such as sand, gypsum, and mica are selected, layered in a specific ratio, and compacted to create a plane stress physical similarity model with a length of 4.0 meters, a height of 3.0 meters, and a width of 0.4 meters. The model simulates a coal seam thickness of 0.025 meters.

[0021] Based on the actual mining area design of the target coal mine, a region with a length of 2.0 meters and a width of 0.3 meters was defined in the physical similarity model as the initial simulated mining area. Using a micro excavator, the first simulated roadway was excavated along the coal seam strike within the initial simulated mining area. The excavation location of the first simulated roadway was 0.5 meters from the model boundary, and the width of the first simulated roadway was 0.03 meters, forming the initial simulated mining face. Within the initial simulated mining area, subsequent simulated roadways were excavated sequentially, parallel to the first simulated roadway and at preset intervals of 0.03 meters. This preset interval is equivalent to the width of the simulated roadway. The separating coal walls between adjacent simulated roadways were removed, connecting the independent simulated roadways to form a continuous space with a uniform width of 0.3 meters. This continuous space is the final simulated mining face. Throughout the excavation process, the excavation timing, spatial coordinates, and the total width of the final simulated mining face (0.3 meters) for each simulated roadway were recorded. In practice, the timing of excavation refers to the excavation of one simulated tunnel at fixed time intervals, from left to right, after the model has stabilized. The time interval is 5 minutes, and a total of 10 simulated tunnels are excavated.

[0022] In some embodiments, the filling mode includes a single-sided filling mode and a double-sided filling mode. In the single-sided filling mode, the simulated filling body is arranged on one side of the simulated mining face along the dip direction; in the double-sided filling mode, the simulated filling body is arranged on both sides of the simulated mining face along the dip direction. A similar material is configured to simulate the actual filling material, and the mechanical strength of the similar material satisfies a similar ratio to the paste filling material planned to be used in the target coal mine. For example, if the 28-day compressive strength of the actual paste filling material is 5 MPa, the target compressive strength of the similar material can be determined based on the strength similarity ratio of the model, and its proportion can be determined through experiments. A feasible proportion is a mixture of water, sand, and cement at a mass ratio of 0.3:1.0:0.15. When the single-sided filling mode is used, a strip-shaped filling body with the same length as the simulated mining face and a width of 0.05 meters is constructed at one boundary of the simulated mining face along the dip direction using the similar material, forming the simulated filling body in the single-sided filling mode. During the construction of the simulated filling body, the thickness and compaction degree of the similar material are controlled to simulate the roof contact effect and density of the actual filling operation. Once the similar material has solidified and stabilized, the simulated filling body is formed. Simultaneously, the specific location, width (0.05 meters), and relative position of the simulated filling body to the simulated mining face are recorded. Optionally, when using a double-sided filling mode, two strip-shaped filling bodies of equal length and 0.025 meters wide are constructed on both sides of the simulated mining face along the dip direction using similar material, forming the simulated filling body under the double-sided filling mode. In this implementation, the total width of the similar material used in the double-sided filling mode is the same as in the single-sided mode, both being 0.05 meters (each filling body is 0.025 meters wide), for comparison of consistent filling material usage. The final mix ratio of the similar material can be determined by experimentally calibrating its relationship with the target strength; the calibration formula can be expressed as:

[0023] in: This indicates the compressive strength that similar materials must achieve. This represents the strength similarity constant derived from the geometric similarity ratio and the bulk density similarity ratio of the model. This represents the compressive strength of the actual filling material. Record the specific location of the simulated filling bodies in the double-sided filling mode, the width of each filling body (0.025 meters), and their relative relationship to the simulated mining face. It can be understood that by implementing the above specific parameters, the differences in spatial arrangement between the single-sided and double-sided filling modes can be clearly compared under the same simulated mining face width and total filling width.

[0024] Referring to Figure 3, continuing with the model based on the actual coal seam thickness of 2.5 meters, mining depth of 300 meters, and geometric similarity ratio of 1:100 in the target coal mine, during the process of completing the simulated mining face excavation and simulated backfill construction, the coal seam area that will not be excavated was clearly planned. In this implementation, the coal seam area that will not be excavated is located between the mineable area of ​​the simulated backfill and the simulated mining face, and is completely preserved in the model, thus forming a simulated coal pillar. The width of the simulated coal pillar is measured and recorded. The width of the simulated coal pillar is the distance from the inner edge of the simulated backfill to the boundary of the simulated mining face along the dip direction of the coal seam. It can be understood that when a single-sided backfill mode is adopted and the width of the simulated backfill is 0.05 meters, if the planned width of the simulated coal pillar is 0.04 meters, then the measured value of the distance from the inner edge of the simulated backfill to the boundary of the simulated mining face is 0.04 meters. It is ensured that the coal seam of the simulated coal pillar maintains its original connection with the roof and floor strata to simulate the support conditions of the coal pillar in an actual mine. When a double-sided filling mode is adopted and the width of the simulated filling body on each side is 0.025 meters, a section of coal seam area that will not be excavated can be planned on each side of the simulated mining face, i.e., two simulated coal pillars, and the width of each simulated coal pillar can be measured and recorded.

[0025] In this implementation, based on the collected subsidence displacement data, a subsidence curve of the roof during the mining process is plotted, and the morphological characteristics of the subsidence curve are analyzed. The specific process is as follows: A two-dimensional coordinate system is established with the strike direction of the simulated mining face as the abscissa and the subsidence displacement of the overlying strata as the ordinate. The subsidence displacement data of multiple displacement monitoring points located in the same stratum and arranged along the strike direction are connected in the order of the monitoring points to plot the roof subsidence curve of the stratum. The overall shape of the plotted roof subsidence curve is observed, and points that are significantly convex relative to the sides are identified as peaks, and points that are significantly concave relative to the sides are identified as troughs. Optionally, the identification process can be carried out by finding points where the first derivative of the subsidence curve is zero and the second derivative is negative as candidate peaks, and points where the first derivative is zero and the second derivative is positive as candidate troughs, and then confirming them in combination with the shape of the subsidence curve. The identified crest locations are compared with the spatial locations of simulated coal pillars in the model, and the identified trough locations are compared with the spatial locations of simulated mining faces in the model. It can be understood that, theoretically, in the strike direction, the crest location should correspond to the location where a simulated coal pillar supports the rock below, while the trough location should correspond to the location where a simulated mining face (already mined out) is located below. The difference between the subsidence displacement values ​​at the crest and trough is calculated; this difference characterizes the unevenness of roof subsidence between the coal pillar support area and the mining face area. The formula for calculating the difference in subsidence displacement between the crest and trough is:

[0026] in: This represents the difference in subsidence displacement between the wave crest and the wave trough. This represents the subsidence displacement at the wave crest. This represents the subsidence displacement value at the trough. For example, in a test, if the subsidence displacement value at a wave crest is identified as 3.2 mm, and the subsidence displacement value at the adjacent wave trough is 8.7 mm, then the calculated subsidence displacement difference is... The value is -5.5 mm. Optionally, the above calculation can be performed on multiple crest-trough pairs along the strike to obtain a series of subsidence displacement differences, which can be used to analyze the spatial distribution characteristics of the roof subsidence unevenness as a whole.

[0027] In this embodiment, the differences in overlying rock movement are analyzed based on the distribution patterns of wave crest and trough positions, and the surface subsidence rate is calculated. This data guides actual mining parameters in the field. The implementation process can be illustrated through the following example scenario and data comparison. The final stable subsidence displacement of all displacement monitoring points deployed on the surface layer is extracted. In this embodiment, it is assumed that 20 displacement monitoring points are uniformly arranged along the strike of the model surface. After one simulated mining operation stabilizes, the subsidence displacement of these 20 monitoring points is recorded as follows: 1.2 mm, 1.5 mm, 1.8 mm, 2.1 mm, 2.3 mm, 2.0 mm, 1.7 mm, 1.9 mm, 2.2 mm, 2.4 mm, 2.6 mm, 2.3 mm, 2.0 mm, 1.8 mm, 1.6 mm, 1.4 mm, 1.3 mm, 1.1 mm, 0.9 mm, and 0.8 mm. The average value of the subsidence displacement of these surface monitoring points is calculated to obtain the maximum surface subsidence value. For example, summing the above 20 values ​​and dividing by 20 yields a maximum surface subsidence value of 1.8 mm. The maximum surface subsidence value can also be the maximum value among all surface monitoring points' subsidence displacements, but the average value better reflects the overall level of surface movement. The simulated coal seam thickness corresponding to the simulated mining operation in the physical similarity model is obtained. Based on the model's geometric similarity ratio of 1:100, if the actual mining thickness is 2.5 meters, the simulated coal seam thickness is 0.025 meters. This can be understood as the surface subsidence rate... The calculation formula is:

[0028] in: Indicates the rate of land subsidence. This represents the maximum subsidence value of the earth's surface. This represents the simulated coal seam thickness. The surface subsidence rate, used to characterize the degree of surface movement, is calculated by dividing the maximum surface subsidence value of 1.8 mm by the simulated coal seam thickness of 25 mm. It is 0.072.

[0029] The surface subsidence rate calculated under the single-sided filling mode was compared with that calculated under the double-sided filling mode. Two sets of control experiments were conducted under the condition that the simulated mining face width and total filling material width were the same. The first set of experiments used the single-sided filling mode, with a simulated filling body width of 0.05 meters and a simulated coal pillar width of 0.04 meters, and the maximum surface subsidence value was measured. The value is 2.1 mm, and the calculated surface subsidence rate for the single-sided model is... The value was 0.084. The second group of tests used a double-sided filling mode, with a simulated filling body width of 0.025 meters on each side and a simulated coal pillar width of 0.02 meters on both sides. The maximum surface subsidence value was measured. The surface subsidence rate of the two-sided model was calculated to be 1.5 mm. The value is 0.060. The surface subsidence rate of 0.084 in the single-sided model and 0.060 in the double-sided model, along with their corresponding simulated mining face width of 0.3 meters and simulated coal pillar widths (0.04 meters in the single-sided model and 0.02 meters * 2 in the double-sided model), are recorded as a set of correlated data. The simulated mining face width or simulated coal pillar width can be changed, and the above experiment and calculation can be repeated to obtain multiple sets of correlated data.

[0030] In this implementation, the actual mining parameters are guided by the surface subsidence rate. The process includes cutting the physical similarity model after the mining operation is completed. The physical similarity model is cut along a profile line perpendicular to the simulated mining face, exposing the overburden profile inside the model. On the exposed profile, observations are made upwards from the top of the coal seam to identify areas where the rock strata have fractured, collapsed, and lost their original layered structure. The height of this area is the height of the caving zone. For example, if the rock strata observed to be broken and randomly piled up within a 0.12-meter height range from the top of the coal seam, the measured height of the caving zone is 0.12 meters. Above the caving zone, observations continue to identify areas with obvious delamination and fissures, but where the rock strata have not completely collapsed. The height of this area is the height of the fracture zone. For example, if numerous vertical fissures and delaminations are observed within a 0.45-meter height range from the top of the caving zone, the measured height of the fracture zone is 0.45 meters. It is understandable that the measured height of the fracture zone (0.45 meters) includes the height of the caving zone (0.12 meters). Based on the observed heights of the caving zone (0.12 meters), the fracture zone (0.45 meters), and the simulated mining face width (0.3 meters), the correlation between the mining face width and the overburden failure height is analyzed. Multiple sets of experiments with different simulated mining face widths are conducted to obtain multiple data pairs of caving zone height, fracture zone height, and mining face width. Combining the analysis results of surface subsidence rate, overburden failure height, and mining face width, the parameter combination for guiding actual on-site mining—including the excavation width, backfill arrangement, and coal pillar width—is determined. For example, if the analysis shows that when the simulated mining face width is 0.25 meters and a double-sided backfilling mode is used, both the surface subsidence rate and the fracture zone height are less than the preset control targets, then this parameter combination (mining face width 0.25 meters, double-sided backfilling) can be used as the recommended mining control parameters.

[0031] In this embodiment, after the physical similarity model mining operation is completed, the model is sectioned to observe and measure the extent of overburden damage. This process can be illustrated through the following example scenario and data comparison. Once all displacement monitoring data within the physical similarity model has stabilized, data acquisition is stopped, and preparation is made for destructive sectioning of the physical similarity model. The physical similarity model is cut along a section line perpendicular to the simulated mining face, exposing the cross-section of the overburden inside the model. On the exposed cross-section, observations are made upwards from the top of the mined coal seam to identify areas where the rock strata have fractured, collapsed, and lost their original layered structure. The height of this area is the height of the collapse zone. For example, through observation and measurement, if the rock strata within a 0.12-meter height range from the top of the coal seam are completely broken and randomly piled up, the measured height of the collapse zone is 0.12 meters. Above the collapse zone, observations continue to identify areas with obvious delamination and fissures, but where the rock strata have not yet completely collapsed. The height of this area is the height of the fissure zone. For example, if numerous vertical fractures and delaminations are observed in the rock strata extending upwards from the top interface of the caving zone (i.e., at 0.12 meters) to a height of 0.45 meters, but the strata still maintain a certain stratification, then the measured height of the fracture zone is 0.45 meters. Using measuring tools, the vertical distance from the top of the mined coal seam to the top interface of the caving zone is directly measured to obtain the measured height of the caving zone, and the vertical distance from the top of the mined coal seam to the top interface of the fracture zone is also measured to obtain the measured height of the fracture zone. It can be understood that the measured height of the fracture zone of 0.45 meters includes the measured height of the caving zone of 0.12 meters.

[0032] Based on the observed heights of the caving zone and fracture zone, as well as the width of the simulated mining face, the correlation between the mining face width and the overlying strata failure height was analyzed. Data pairs were established between the measured heights of the caving zone obtained under different experimental schemes and the corresponding simulated mining face widths. For example, five sets of data pairs were obtained by conducting five sets of experiments with different simulated mining face widths, as shown in Table 1.

[0033] Table 1: Measured Values ​​of Overburden Failure Height under Different Simulated Mining Face Widths Test Scheme No. Simulated Mining Face Width Height of Collapse Zone Measured Value of Fracture Zone Height Measured Value 10.20 0.08 0.30 20.25 0.10 0.38 30.30 0.12 0.45 40.35 0.15 0.60 50.40 0.18 0.75 Table 1 analyzes the trend of the caving zone height as the simulated mining face width changes. As can be seen from the data in Table 1, as the simulated mining face width W increases from 0.20 meters to 0.40 meters, the measured height H of the caving zone... cThe height of the fracture zone increased from 0.08 meters to 0.18 meters, showing an increasing trend. Data pairs were established between the measured values ​​of the fracture zone height obtained under different experimental schemes and the corresponding simulated mining face widths to analyze the trend of fracture zone height variation with simulated mining face width. As shown in Table 1, with the increase of the simulated mining face width W, the measured value of the fracture zone height H... f The diameter increased from 0.30 meters to 0.75 meters, also showing an increasing trend.

[0034] In practical implementation, a fracture zone height development threshold is set, and test schemes where the measured fracture zone height is less than the threshold are identified. The simulated mining face width conditions corresponding to these test schemes are recorded. Assuming the fracture zone height development threshold is set to 0.50 meters according to engineering requirements, Table 1 identifies that the measured fracture zone heights (0.30 meters, 0.38 meters, and 0.45 meters) for test schemes 1, 2, and 3 are all less than 0.50 meters. The simulated mining face width conditions corresponding to these schemes are recorded as 0.20 meters, 0.25 meters, and 0.30 meters, respectively. It can be understood that the fracture zone height development threshold is a preset boundary value used to evaluate whether overburden failure is controllable. The change in the fracture zone height growth rate during the gradual increase of the simulated mining face width is analyzed. The fracture zone height growth rate R can be calculated from adjacent data points using the following formula:

[0035] in: Indicates the height growth rate of the fracture zone. This represents the change in the measured height of the fracture zone between two adjacent tests. This indicates the change in the width of the simulated mining face between two consecutive tests. For example, according to the data in Table 1, when the width of the simulated mining face increases from 0.30 meters (Scheme 3) to 0.35 meters (Scheme 4), rice, The calculated height growth rate R of the fracture zone in this interval is 3.0 (m / m). Optionally, the growth rate for each width interval is calculated, and the change in the R value is analyzed. The critical value of the simulated mining face width corresponding to the slowdown in the height growth rate of the fracture zone is determined. For example, if the calculation shows that when the simulated mining face width is less than 0.30 meters, the average growth rate R1 is 1.6 (m / m); and when the simulated mining face width is greater than 0.30 meters, the average growth rate R2 is 3.0 (m / m), then the growth rate has not slowed down but accelerated. In this case, it may be necessary to pay attention to the inflection point where the growth rate begins to increase significantly. For example, a simulated mining face width of 0.30 meters can be considered a critical width; beyond this width, fracture zone development significantly intensifies.

[0036] Referring to Figure 4, in the physical similarity simulation test of the coupled mining control method of coal mine excavation and filling, the response law of surface subsidence rate with the change of simulated mining face width under different filling modes can be quantitatively analyzed by this curve. In Figure 4, the horizontal axis is the simulated mining face width (m), covering the test range of 0.20m to 0.40m; the vertical axis is the surface subsidence rate (dimensionless), representing the ratio of the maximum surface subsidence value to the simulated mining thickness, used to quantify the degree of surface movement. The curve data shows that as the simulated mining face width gradually increases from 0.20m to 0.40m, the surface subsidence rate under both filling modes shows a monotonically increasing trend, but the growth rate is significantly different: single-sided filling mode (dotted curve): the surface subsidence rate continuously increases from 0.15 to 0.35, which is higher than that of the double-sided filling mode under all width conditions, indicating that the control ability of the single-sided support structure on the overburden and surface is relatively weak. Double-sided filling mode (boxed curve): The surface subsidence rate gradually increased from 0.10 to 0.22, with a more gradual overall increase, demonstrating that the double-sided symmetrical support structure can more effectively constrain overburden movement and reduce the risk of surface subsidence. Analysis of the curve slope shows that when the simulated mining face width exceeds 0.30m, the surface subsidence rate increases significantly under the single-sided filling mode, while the double-sided filling mode maintains a relatively stable growth trend. This phenomenon indicates that under wide mining face conditions, the double-sided filling mode has a more prominent advantage and can be used as a priority scheme for controlling surface movement. The experimental results provide direct evidence for optimizing field mining parameters: under the premise of meeting the safety threshold for fracture zone development height, the double-sided filling mode should be prioritized, and the parameters for coal pillar retention and filling body layout should be rationally determined in conjunction with the critical mining face width to achieve the dual objectives of overburden stability and surface subsidence control.

[0037] Based on the calculated surface subsidence rate, a filling mode that meets the surface movement control target is selected. For example, assuming the surface movement control target requires a surface subsidence rate of less than 0.065, in the simulation test, the surface subsidence rate corresponding to the single-sided filling mode is 0.084, and the surface subsidence rate corresponding to the double-sided filling mode is 0.060. Therefore, the double-sided filling mode meets the control target and is selected as the recommended filling mode for the field. Based on the correlation analysis results between the measured height of the fracture zone and the simulated mining face width, the maximum allowable mining face width in actual field mining is determined. In this embodiment, the safety threshold for fracture zone height development is set at 0.50 meters. As shown in Table 1, when the simulated mining face width is 0.30 meters, the fracture zone height is 0.45 meters, which is less than 0.50 meters; when the simulated mining face width is 0.35 meters, the fracture zone height is 0.60 meters, which exceeds the safety threshold. Therefore, the maximum mining face width is determined to be 0.30 meters. This width should ensure that the fracture zone development height is controlled below the preset safety threshold of 0.50 meters. It is understood that this maximum mining face width is a simulated value obtained from physical similarity model tests.

[0038] In practical implementation, the measured height of the caving zone and the ratio of the mining thickness are used to verify whether the pre-set strength of the backfill can effectively support the weight of the rock strata within the caving zone. For example, in one test, the measured height of the caving zone was 0.12 meters, and the simulated mining thickness in the model was 0.025 meters, resulting in a caving zone height to mining thickness ratio of 4.8. Assuming that this ratio corresponds to a caving zone height of approximately 12 meters under field conditions through similarity theory conversion, the strength design of the backfill needs to be checked based on the weight of the rock strata within this caving zone height to verify its supporting capacity. Considering the experimental phenomenon that the presence of simulated backfill can reduce the width of the coal pillar, and under the premise of ensuring stable support from both the simulated coal pillar and the simulated backfill, an empirical relationship between the coal pillar width, the backfill width, and the mining face width is established through fitting multiple sets of experimental data. Based on multiple sets of experimental data using a double-sided backfilling mode, a linear empirical relationship is obtained through fitting:

[0039] in: This indicates the width of a single coal pillar in a double-sided filling mode. Indicates the width of the simulated mining face. This indicates the width of the simulated filling material on one side when filling on both sides. and These are coefficients obtained through data fitting. This represents the weighting of the influence of the simulated filling width on the width of a single-sided coal pillar. For example, if coefficient a is 0.1, coefficient c is 0.01, and coefficient b is 0, based on fitting multiple sets of experimental data, then the empirical relationship is: .

[0040] In practice, based on empirical formulas, the maximum mining face width, and the selected backfilling mode, the recommended actual coal pillar width and backfilling arrangement width are calculated. Actual geological conditions on-site, including actual mining depth, actual mining thickness, and lithological strength, are converted into corresponding simulation parameters according to the similarity ratio followed by the physical similarity model. Assuming the actual planned mining face width is 30 meters, based on the model's geometric similarity ratio of 1:100, the simulated mining face width in the model is converted to 0.30 meters, which is equal to the previously determined maximum mining face width. The mined face width in the converted simulation parameters is set to the maximum mining face width of 0.30 meters and substituted into the empirical formula corresponding to the backfilling mode (double-sided backfilling) selected based on experimental data. In the empirical formula, the width of the filling material is used as an input variable to solve for the calculated value of the coal pillar width required to achieve stable support. The calculation process is as follows: Substituting B = 0.30 meters into the formula, we obtain... Meters. This calculated value. The meter represents the width of a single-sided coal pillar at the model scale. The calculated value of the coal pillar width is compared with the minimum safe coal pillar width calculated based on the theory of in-situ rock strata strength. Assuming that the minimum safe coal pillar width (single-sided) required under the same conditions, calculated according to the in-situ rock strata strength theory, is 0.035 meters when converted to the model scale, then the larger of the calculated coal pillar width (0.04 meters) and the minimum safe coal pillar width (0.035 meters), i.e., 0.04 meters, is taken as the final recommended value for the actual coal pillar width at the model scale. Based on the finally determined actual coal pillar width (model scale 0.04 meters), the maximum mining face width (model scale 0.30 meters), and empirical formulas, the corresponding actual backfill width is calculated. If the backfill width in the empirical formula is a fixed value, the experimental value is directly used. For example, in the double-sided backfill mode, the width of the single-sided backfill in the model is 0.025 meters. Finally, based on the geometric similarity ratio of 1:100 of the model, the parameters at the model scale were converted back to the actual field scale: the recommended actual field mining face width is 30 meters, the actual coal pillar width on one side is 4 meters, and the actual filling body layout width on one side is 2.5 meters, thus completing the determination of mining control parameters.

[0041] Referring to Figure 5, in the physical similarity simulation experiment, the quantitative response relationship between the simulated mining face width and the overburden failure height was revealed through cross-sectional observation of the model. Specifically, as the simulated mining face width gradually increased from 0.20m to 0.40m, the development heights of both the caving zone and the fracture zone within the overburden showed a monotonically increasing trend: the caving zone height (dotted curve) increased relatively slowly, rising gradually from 0.08m to 0.18m, indicating that its development was less sensitive to the mining face width; the fracture zone height (boxed curve) increased significantly, linearly increasing from 0.30m to 0.75m, indicating a strong positive correlation between its development height and the mining face width. The safety threshold for the fracture zone was set at 0.50m (dashed line). When the simulated mining face width reached 0.30m, the fracture zone height was 0.45m, still within a safe and controllable range; however, when the width exceeded 0.30m (e.g., 0.35m), the fracture zone height (0.60m) exceeded the safety threshold. Therefore, the 0.30m indicated by the dotted dashed line in the figure represents the maximum allowable width of the mining face where the fracture zone development is controlled within the safety threshold. This relationship curve provides a direct basis for optimizing on-site mining parameters: in the double-sided backfilling mode, to control the fracture zone height below the safety threshold, the simulated mining face width should not exceed 0.30m. This value, after conversion by similarity ratio, can guide the determination of the actual mining face width on-site.

[0042] The above are merely preferred embodiments of the present invention and are not intended to limit the present invention in any other way. Any person skilled in the art may make changes or modifications to the above-disclosed technical content to create equivalent embodiments that can be applied to other fields. However, any simple modifications, equivalent changes, and modifications made to the above embodiments based on the technical essence of the present invention without departing from the scope of the present invention shall still fall within the protection scope of the present invention.

Claims

1. A mining control method based on the coupling of coal mine excavation, mining, filling, and retention, characterized in that, The method includes: constructing a physical similarity model for simulating the geological conditions of a target coal mine; excavating simulated roadways in the physical similarity model according to a preset tunneling sequence to form a simulated mining face for simulating longwall mining operations, and recording the initial width and spatial position of the simulated mining face; performing simulated filling operations within or beside the simulated mining face according to a preset filling pattern to form a simulated filling body; retaining a simulated coal pillar between the simulated mining face and the simulated filling body, so that the simulated filling body and the simulated coal pillar are spatially adjacent to each other to form a support structure; and in the physical similarity model... During mining and backfilling operations, multiple sets of displacement monitoring points are deployed to collect real-time subsidence and displacement data of each overburden stratum and surface in the physical similarity model. Based on the collected subsidence and displacement data, a subsidence curve of the roof during mining is plotted, and the morphological characteristics of the subsidence curve are analyzed to identify the peak and trough positions. Based on the distribution patterns of the peak and trough positions, the difference in movement of the overburden at corresponding positions of the simulated coal pillar and the simulated mining face is analyzed, and the ratio of surface subsidence to simulated mining thickness is calculated to obtain the surface subsidence rate. Based on the surface subsidence rate, actual mining parameters are guided on site.

2. The mining control method based on the coupling of coal mine excavation, mining, filling, and retention according to claim 1, characterized in that, In the physical similarity model, simulated roadways are excavated according to a preset tunneling sequence to form a simulated mining face for simulating mining operations. This includes: the physical similarity model is scaled down based on the actual borehole columnar section, rock mechanics parameters, and mining depth data of the target coal mine; an initial simulated mining area is delineated in the physical similarity model according to the actual mining area design of the target coal mine; a micro excavation tool is used to excavate the first simulated roadway along the coal seam strike within the initial simulated mining area to form the initial simulated mining face; within the initial simulated mining area, subsequent simulated roadways are excavated sequentially, parallel to the first simulated roadway and at preset intervals; the separating coal walls between adjacent simulated roadways are removed, connecting the independent simulated roadways to form a continuous space with a uniform width, which is the final simulated mining face; throughout the excavation process, the excavation timing, spatial coordinates, and total width of the final simulated mining face of each simulated roadway are recorded.

3. The mining control method based on the coupling of coal mine excavation, mining, filling, and retention according to claim 2, characterized in that, The filling mode includes a single-sided filling mode and a double-sided filling mode; in the single-sided filling mode, the simulated filling body is arranged on one side of the simulated mining face; in the double-sided filling mode, the simulated filling body is arranged on both sides of the simulated mining face; a similar material is configured to simulate the actual filling material, and the mechanical strength of the similar material satisfies a similar proportional relationship with the actual filling material planned to be used in the target coal mine; when the single-sided filling mode is adopted, at the boundary of the simulated mining face along the dip side, a strip-shaped filling body of the same length and a set width as the simulated mining face is constructed using the similar material to form the... A simulated backfill body in a single-sided backfilling mode; when a double-sided backfilling mode is used, strip-shaped backfill bodies of the same length and a set width as the simulated mining face are constructed at the two sides of the dip of the simulated mining face using the same material, forming the simulated backfill body in the double-sided backfilling mode; during the construction of the simulated backfill body, the laying thickness and compaction degree of the same material are controlled to simulate the roof contact effect and density of the actual backfilling operation; after the same material has solidified and stabilized, the simulated backfill body is formed, and the specific position, width and relative relationship of the simulated backfill body to the simulated mining face are recorded.

4. The mining control method based on the coupling of coal mine excavation, mining, filling, and retention according to claim 3, characterized in that, During the excavation process to form the simulated mining face and construct the simulated backfill, coal seam areas that will not be excavated are clearly defined. These coal seam areas are located between the mineable areas of the simulated backfill and the simulated mining face. The coal seam areas are completely preserved in the model, forming the simulated coal pillar. The width of the simulated coal pillar is measured and recorded. The width of the simulated coal pillar is the distance from the inner edge of the simulated backfill to the boundary of the simulated mining face along the dip direction of the coal seam. Ensure that the coal seam of the simulated coal pillar maintains its original connection with the roof and floor strata to simulate the support conditions of an actual coal pillar.

5. A mining control method based on the coupling of coal mine excavation, mining, filling, and retention according to claim 4, characterized in that, Based on the collected subsidence displacement data, a subsidence curve of the roof during mining is plotted, and the morphological characteristics of the subsidence curve are analyzed, including: establishing a two-dimensional coordinate system with the strike direction of the simulated mining face as the abscissa and the subsidence displacement of the overburden as the ordinate; connecting the subsidence displacement data of multiple displacement monitoring points located in the same rock stratum and arranged along the strike direction according to the order of the monitoring point positions to plot the roof subsidence curve of the rock stratum; observing the overall shape of the roof subsidence curve, identifying points that are significantly convex relative to the two sides as peaks, and identifying points that are significantly concave relative to the two sides as troughs; comparing the identified peak positions with the spatial positions of the simulated coal pillars in the physical similarity model, and comparing the identified trough positions with the spatial positions of the simulated mining face in the physical similarity model; calculating the difference between the subsidence displacement value at the peak and the subsidence displacement value at the trough, the difference being used to characterize the unevenness of roof subsidence between the coal pillar-supported and mining face areas.

6. The mining control method based on the coupling of coal mine excavation, mining, filling, and retention according to claim 5, characterized in that, Based on the distribution patterns of the peak and trough positions, the differences in the movement of overburden at the corresponding positions of the simulated coal pillar and the simulated mining face are analyzed, and the surface subsidence rate is calculated. This includes: extracting the final stable subsidence displacement of the displacement monitoring points arranged on the surface layer, calculating the average value of the subsidence displacement of all surface monitoring points, and obtaining the maximum surface subsidence value; obtaining the simulated coal seam thickness value of the simulated mining operation in the physical similarity model; dividing the maximum surface subsidence value by the simulated coal seam thickness value to calculate the surface subsidence rate used to characterize the degree of surface movement; comparing the surface subsidence rate of the single-sided model calculated in the single-sided filling mode with the surface subsidence rate of the double-sided model calculated in the double-sided filling mode; and recording the surface subsidence rate of the single-sided model, the surface subsidence rate of the double-sided model, and their corresponding simulated mining face width and simulated coal pillar width as a related data group.

7. A mining control method based on the coupling of coal mine excavation, mining, filling, and retention according to claim 6, characterized in that, Based on the surface subsidence rate, the following parameters guide actual on-site mining: After the mining operation of the physical similarity model is completed, the physical similarity model is sectioned to observe and measure the damage range of the overburden, and to distinguish the height of the caving zone and the height of the fracture zone; based on the observed height of the caving zone, the height of the fracture zone, and the width of the simulated mining face, the correlation between the width of the mining face and the height of the overburden damage is analyzed; combining the analysis results of the surface subsidence rate, the height of the overburden damage, and the width of the mining face, the combination of parameters for guiding actual on-site mining, including the tunneling width, the arrangement of the backfill body, and the width of the coal pillar, is determined.

8. The mining control method based on the coupling of coal mine excavation, mining, filling, and retention according to claim 7, characterized in that, After the mining operation of the physical similarity model is completed, the physical similarity model is sectioned to observe and measure the extent of overburden damage. This includes: stopping data acquisition after all displacement monitoring data within the physical similarity model stabilizes, and preparing to perform destructive sectioning on the physical similarity model; cutting the physical similarity model along a profile line perpendicular to the simulated mining face to expose the profile of the overburden inside the physical similarity model; observing upwards from the top of the mined coal seam on the exposed profile to identify areas where the rock strata have fractured, collapsed, and lost their layered structure, and the height of these areas is the height of the collapse zone; continuing to observe and identify areas above the collapse zone where there are obvious delaminations and fissures but the rock strata have not completely collapsed, and the height of these areas is the height of the fracture zone; using measuring tools, directly measuring the vertical distance from the top of the mined coal seam to the top interface of the collapse zone to obtain the measured value of the height of the collapse zone, and measuring the vertical distance from the top of the mined coal seam to the top interface of the fracture zone to obtain the measured value of the height of the fracture zone.

9. A mining control method based on the coupling of coal mine excavation, mining, filling, and retention according to claim 8, characterized in that, Based on the observed heights of the collapse zone, the fracture zone, and the simulated mining face, the correlation between the mining face width and the overlying strata failure height is analyzed. This includes: establishing data pairs between the measured heights of the collapse zone obtained under different test schemes and the corresponding simulated mining face widths, and analyzing the trend of the collapse zone height changing with the simulated mining face width; establishing data pairs between the measured heights of the fracture zone obtained under different test schemes and the corresponding simulated mining face widths, and analyzing the trend of the fracture zone height changing with the simulated mining face width; setting a fracture zone height development threshold, identifying test schemes where the measured height of the fracture zone is less than the fracture zone height development threshold, and recording the simulated mining face width conditions corresponding to the test schemes; analyzing the change in the fracture zone height growth rate as the simulated mining face width gradually increases, and determining the critical value of the simulated mining face width corresponding to the slowdown in the fracture zone height growth rate.

10. A mining control method based on the coupling of coal mine excavation, mining, filling, and retention according to claim 9, characterized in that, The method also includes steps for optimizing coal pillar and backfill parameters based on simulation test results: Based on the calculated surface subsidence rate, a backfill mode that meets the surface movement control target is selected, which can be either a single-sided or double-sided backfill mode; based on the correlation analysis between the measured height of the fracture zone and the width of the simulated mining face, the maximum allowable mining face width in actual field mining is determined, ensuring that the fracture zone development height is controlled below a preset safety threshold; based on the ratio of the measured height of the caving zone to the mining thickness, the preset backfill strength is verified to effectively support the weight of the rock strata within the caving zone; considering the experimental phenomenon that the presence of the simulated backfill can reduce the width of the coal pillar, and under the premise of ensuring the stability of the simulated coal pillar and simulated backfill, an empirical relationship between the coal pillar width, backfill width, and mining face width is established through fitting multiple sets of experimental data; based on the empirical relationship, the maximum mining face width, and the selected backfill mode, calculations are performed... The recommended actual coal pillar width and backfill layout width are determined by: converting actual geological condition parameters, including actual mining depth, actual mining thickness, and lithological strength, into simulation parameters corresponding to the physical similarity model according to a similarity ratio; setting the mining face width in the converted simulation parameters as the maximum mining face width and substituting it into the empirical formula corresponding to the selected backfill mode; using the backfill layout width as an input variable in the empirical formula to solve for the calculated value of the coal pillar width required to achieve stable support; comparing the calculated value of the coal pillar width with the minimum safe coal pillar width calculated based on the theory of on-site rock strata strength, and taking the larger of the two values ​​as the final recommended actual coal pillar width; and calculating the matching actual backfill layout width based on the final determined actual coal pillar width, the maximum mining face width, and the empirical formula, thus completing the determination of mining control parameters.

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