End coal caving area surrounding rock stability control method and device based on numerical simulation

Abstract

The invention discloses an end coal caving area surrounding rock stability control method and device based on numerical simulation, and belongs to the technical field of surrounding rock control, and the method comprises the steps: S1, analyzing a bearing structure of roadway surrounding rock in an end coal caving area; s2, establishing an end coal caving numerical simulation model by adopting PFC (Power Factor Correction) software; s3, performing analysis based on a simulation result of the end coal caving numerical simulation model in the S2, wherein the analysis comprises active support space-time performance analysis, end coal caving area coal gangue flow form analysis and roadway surrounding rock bearing structure bearing capacity evolution analysis; according to the end coal caving area surrounding rock stability control method and device based on numerical simulation, the coal rock flowing rule, the end surrounding rock stability and the reasonable end coal caving range of end coal rock in the caving process are studied, the safety and efficiency of coal mining are improved, the cost is reduced, and technical innovation is promoted.

Description

Technical Field

[0001] The present invention relates to the technical field of surrounding rock control, and in particular to a method and device for controlling the stability of surrounding rock in the end caving area based on numerical simulation. Background Technique

[0002] Fully-mechanized caving mining is a revolutionary technology for mining thick coal seams, which solves the problems such as stress concentration caused by the remaining coal pillars in the upper slice during the slicing mining of thick coal seams, easy spontaneous combustion in the goaf of the lower slice, difficult roadway support, large relative gas emission in the first mining slice, low output, and high cost. Compared with the large mining height mining technology, fully-mechanized caving mining has the advantages of less investment, low cost, low energy consumption, low emissions, and large adaptability to coal seam thickness changes. However, how to improve the recovery rate of top coal in the fully-mechanized caving mining face has become one of the urgent problems to be solved in the further development of fully-mechanized caving mining technology.

[0003] To reduce top coal loss and improve the top coal recovery rate, many scholars have carried out relevant research from aspects such as the top coal fragmentation mechanism and discharge law, optimization of caving technology, and innovation of caving methods, and achieved remarkable results. However, the research areas in the existing results mainly focus on the middle position of the working face. Due to space limitations in the end area and considering the roadway stability, generally no caving or a small amount of caving is carried out, so the attention is relatively low. However, relevant research shows that the un-discharged top coal above the transition support and end support accounts for 35%-45% of the total coal loss in the fully-mechanized caving face, which has long restricted the improvement of the top coal recovery rate. And there is less research on the coal and rock flow law during the caving process of end coal and rock, the stability of end surrounding rock, and the reasonable end caving range. Therefore, it is necessary to further study. Summary of the Invention

[0004] The purpose of the present invention is to provide a method and device for controlling the stability of surrounding rock in the end caving area based on numerical simulation to solve the problems existing in the above background technique.

[0005] To achieve the above purpose, the present invention provides a method for controlling the stability of surrounding rock in the end caving area based on numerical simulation, including the following steps:

[0006] S1. Analyze the bearing structure of the roadway surrounding rock in the end caving area;

[0007] S2. Establish a numerical simulation model for end caving using PFC software;

[0008] S3. Analyze based on the simulation results of the numerical simulation model for end caving in S2, including the spatio-temporal analysis of active support, the analysis of the coal and gangue flow pattern in the end caving area, and the analysis of the evolution of the bearing capacity of the roadway surrounding rock bearing structure.

[0009] Preferably, step S1 is specifically:

[0010] Before the caving of broken coal and rock mass in the end area of the fully-mechanized caving face, due to the lack of movement space, the coal and rock mass are closely interlocked with each other. As the caving progresses, the originally closely interlocked coal and rock mass around the roadway gradually loosens with the caving of the top coal. When the loosening reaches a certain degree, the load-bearing structure formed by the close interlocking of the coal and rock mass around the roadway will become unstable, increasing the difficulty of passive support in the roadway. If the passive support capacity in the roadway is insufficient, the roadway is at risk of instability, leading to serious safety accidents. Therefore, it is very necessary to ensure the stability of the surrounding rock in the end area. When caving coal at the end, the coal and rock mass in the end area of the working face have been broken into loose blocks. Combining with the Prandtl equilibrium theory, at this time, due to the redistribution of the surrounding rock stress, a pressure arch in the shape of a parabola will be formed above the roadway, which is the equilibrium arch; and along the direction with an angle of with the horizontal direction, two slip surfaces will be formed; the range formed by the equilibrium arch and the slip surfaces is the range where the surrounding rock is damaged;

[0011] The height of the natural equilibrium arch is expressed as:

[0012]

[0013] where, σ v is the vertical pressure at the top of the roadway; γ is the unit weight of the surrounding rock; h is the height of the equilibrium arch; B is the semi-span of the equilibrium arch; b is the semi-span of the roadway; h 0 is the height of the roadway; is the calculated friction angle of the surrounding rock; f is the Prandtl surrounding rock strength coefficient, as follows:

[0014]

[0015] where, R c is the uniaxial saturated compressive strength of the surrounding rock; c is the cohesion of the surrounding rock;

[0016] Before the loose coal and rock mass in the natural equilibrium arch are caved, the rock mass around the ends of the bolts or cables is still closely interlocked with each other, and the active support effect will not fail. Moreover, under the reinforcement effect of the pre-tightening force of the bolts or cables, the broken coal and rock mass in the equilibrium arch will form a stable load-bearing structure within the direct action range of the bolt support, which is the load-bearing arch; the essence of the formation of the load-bearing arch is the active support effect generated by the pre-tightening force applied by the bolts, combined with the passive support to expand the support range, thereby improving the stability of the inlaid composite force chain structure between the fractured rock masses.

[0017] Based on the Mohr-Coulomb criterion, the frictional shear force of the fractured rock mass is expressed as:

[0018]

[0019] where, σ n is the normal stress between the rock masses.

[0020] The load-bearing arch has a load-bearing effect due to the biting force F formed by the mutual extrusion of the crushed stones. i It is sufficient to balance the self-weight G of the crushed stones, and the calculation formula is:

[0021] F i = F l μ (4)

[0022] Among them, F l is the lateral binding force on the crushed stones at the arch crown; μ is the equivalent friction coefficient between the crushed stones at the arch crown;

[0023] Only when the crushed stones at the arch feet of the load-bearing arch become loose and the biting force between the crushed stones is difficult to bear the self-weight, will the load-bearing arch be damaged and the roadway lose its load-bearing capacity. Therefore, when performing end caving, the stability of the load-bearing arch is used to ensure the safety of the roadway. The formation range of the load-bearing arch is within the direct action range of the bolt support, which is:

[0024] h p = αβL (5)

[0025] Among them, h p is the height of the load-bearing arch; α is the correction coefficient of the load-bearing arch height, and 0 < α < 1; β is the safety factor, 1 ≤ β ≤ 1 / α; L is the length of the direct action range of the bolt.

[0026] Preferably, during end caving in S2, the upper and lower ends of the working face are far apart and will not affect each other. Models are established for the upper and lower end regions of the working face respectively; the roadway support combines active support and passive support. Considering that the actual generation of bolts or cables in two-dimensional simulation will produce a limited small boundary, therefore, according to the actual bolt row spacing, the pre-tightening force is applied by using the geometry module for equivalent substitution. The bolts or cables in the roadway are numbered as the middle cable mc, and the right part of the roadway is numbered clockwise from the middle cable as r 1 ~r 7 , and the left half of the roadway is numbered counterclockwise from the middle cable as l 1 ~l 7 ; The passive support uses Fish language to generate wall elements to simulate the shed.

[0027] Preferably, the initial state of the model in S2 includes six parts: the 82 coal seam, the parting layer, the 81 coal seam, the immediate roof, the main roof, and the overlying rock. Among them, the density of the overlying rock is converted according to the following formula using Fish language:

[0028]

[0029] Among them, P is the load of the overlying strata; g is the acceleration due to gravity; h is the height in the rectangular area of the overlying rock in the two-dimensional model.

[0030] Preferably, the spatio-temporal analysis of active support in S3 is to analyze the spatio-temporal differential characteristics of the active support effect in the roadway. During the coal caving process, taking a basic time unit, respectively count the effective periods of active support during the coal caving process at the upper and lower ends of the working face with different dips, including:

[0031] Lower end: During the initial coal caving process at the lower end of the working face with different dips, although there are differences in the time nodes when the bolts or cable bolts in the roadway fail, after the coal caving ends, the active support effects of the bolts or cable bolts in the middle and right parts of the roadway are basically invalidated. Strengthen the passive support in the middle and right parts of the roadway before coal caving to ensure that the load-bearing arch around the roadway can still maintain its load-bearing capacity after the active support fails;

[0032] Upper end: When the coal caving reaches the position of the transition support, the active support effects of the bolts or cable bolts in the middle and left parts of the roadway are basically invalidated. Before the coal caving at the upper end of the working face, strengthen the passive support in the roadway to ensure the stability of the arch feet of the load-bearing arch.

[0033] Preferably, the analysis of the flow pattern of coal and gangue in the coal caving area at the end of S3 is to observe and evaluate the flow and migration characteristics of the coal and rock mass and the evolution law of the surrounding rock structure of the roadway during the coal caving process at the end of the face, taking a basic time unit, including:

[0034] Lower end: When coal is caved at the same position on the working face with different dips, the failure speed of the load-bearing arch at the lower end of the working face is negatively correlated with the dip of the working face;

[0035] Upper end: The coal caving in the middle of the working face has an advanced influence on the loosening of the load-bearing arch at the upper end, and the advanced influence distance is positively correlated with the dip of the working face.

[0036] Preferably, the analysis of the evolution of the load-bearing capacity of the surrounding rock load-bearing structure in S3 is to analyze the force chain field during the coal caving process to show the dynamic change process of the load-bearing capacity of the load-bearing arch. The force chain is the main form of non-continuous and granular medium to transfer loads. Through the force chain division method, the force chain is divided into three grades: strong force chain, secondary strong force chain and weak force chain according to the strength, as:

[0037]

[0038] Among them, F is the contact force between particles; is the average contact force;

[0039] The combined force chain structure formed by the three strength force chains breaks and regenerates continuously throughout the coal caving process, including the connection of strong force chains - the break of strong force chains - the reconstruction of strong force chains. In the model, the Fish language is used to screen and distinguish different strength force chains, and analyze the change process of the force chain field during the coal caving process at the end of the face, including:

[0040] Lower end head: The bearing capacity of the bearing arch at the lower end of the working face is positively correlated with the dip angle of the working face;

[0041] Upper end head: The bearing capacity of the bearing arch at the upper end of the working face is negatively correlated with the dip angle of the working face.

[0042] A device for a method of controlling the stability of surrounding rocks in the end caving area based on numerical simulation, for the caving requirements in the end caving area, including an end caving support. The end caving support includes a shield beam. There is a window at the top of the shield beam. Hydraulic cylinders are respectively arranged on both sides of the shield beam. A flexible metal net is arranged in the window. The extension and retraction of the flexible metal net are controlled by the hydraulic cylinders to achieve precise control of the coal discharge port and complete the end caving operation. The control system of the hydraulic cylinders is integrated into the intelligent electro-hydraulic control module of the whole device to realize the intelligent fully-mechanized caving of the end caving support, improve the efficiency and accuracy of mining operations. A coal breaking knife is arranged at one end of the flexible metal net. A coal breaking knife groove adapted to the coal breaking knife is arranged on the shield beam. A net winding device is arranged below the shield beam. A net winding guide plate is arranged on one side of the net winding device, and a coal discharge port is arranged on the side of the flexible metal net close to the coal breaking knife.

[0043] Therefore, the present invention adopts the above-mentioned method and device for controlling the stability of surrounding rocks in the end caving area based on numerical simulation to study the coal and rock flow law, the stability of end surrounding rocks and the reasonable end caving range during the caving process of end coal and rock, which not only improves the safety and efficiency of coal mining, but also reduces the cost and promotes technological innovation.

[0044] The technical solution of the present invention will be further described in detail below through the drawings and embodiments. Brief Description of the Drawings

[0045] Figure 1 It is a flowchart of the method for controlling the stability of surrounding rocks in the end caving area based on numerical simulation of the present invention;

[0046] Figure 2 It is a schematic diagram of the Prandtl equilibrium arch in the embodiment of the present invention;

[0047] Figure 3 It is a schematic diagram of the bearing arch in the embodiment of the present invention;

[0048] Figure 4 It is a schematic diagram of the equivalent bolt in the embodiment of the present invention;

[0049] Figure 5 It is a schematic diagram of the real-time monitoring process in the embodiment of the present invention;

[0050] Figure 6 It is a statistical schematic diagram of the spatio-temporal effect of active support at the lower end head in the embodiment of the present invention;

[0051] Figure 7Schematic diagram for statistical analysis of the spatio-temporal effect of the active support at the lower end of the embodiment of the present invention;

[0052] Figure 8 Schematic diagram of the coal caving displacement field at the lower end with different dip angles in the embodiment of the present invention;

[0053] Figure 9 Schematic diagram of the coal caving displacement field at the upper end with different dip angles in the embodiment of the present invention;

[0054] Figure 10 Schematic diagram of the force chain field of coal caving at the lower end with different dip angles in the embodiment of the present invention;

[0055] Figure 11 Schematic diagram of the zoning of the danger degree of coal caving at the end in the embodiment of the present invention;

[0056] Figure 12 Schematic diagram of the force chain state of coal caving at the upper end with different dip angles in the embodiment of the present invention;

[0057] Figure 13 Schematic diagram of the in-frame segmentation of the No. 106 support in the embodiment of the present invention;

[0058] Figure 14 Schematic diagram of the ultimate coal caving progress of the working face with different dip angles in the embodiment of the present invention;

[0059] Figure 15 Overall schematic diagram of the coal caving support at the end in the embodiment of the present invention;

[0060] Figure 16 Internal schematic of the coal caving support at the end in the embodiment of the present invention Figure 1 ;

[0061] Figure 17 Internal schematic of the coal caving support at the end in the embodiment of the present invention Figure 2 ;

[0062] Reference numerals: 1, shield beam; 2, flexible metal mesh; 3, coal breaking knife; 4, hydraulic cylinder; 5, coal breaking knife groove; 6, net winding device; 7, net winding guide plate; 8, coal discharge opening. Detailed Description of the Invention

[0063] The following detailed description of the embodiments of the present invention provided in the drawings is not intended to limit the scope of the claimed invention, but merely represents selected embodiments of the present invention. All other embodiments obtained by those of ordinary skill in the art based on the embodiments of the present invention without creative efforts fall within the scope of protection of the present invention.

[0064] In this embodiment, taking the coal mine of a certain mining group as an example, the 825 fully-mechanized caving face mines the 82nd coal seam and the 81st coal seam. The average thickness of the 82nd coal seam is 2.50 m, the average thickness of the 81st coal seam is 2.84 m, and there is a parting layer with an average thickness of 1.94 m between the 82nd coal and the 81st coal. The average dip angle of the coal seam is 10°, belonging to a gently inclined coal seam. The working face depth is 262 - 380 m. The strike length of the working face is 681 m, the inclined width is 160 m, a total of 6 transition supports with a center distance of 1.5 m and 102 intermediate supports with a center distance of 1.5 m are arranged, the mining-to-caving ratio is 1:1.91, and the coal caving sequence is from bottom to top. The cross-section of the working face roadway is a straight-wall semi-circular arch, the roadway width is 5 m, and the distance from the arch crown to the roadway floor is 3.75 m. Bolt, cable and timbering combined support is adopted, the bolt length is 2400 mm, the cable length is 6300 mm, and the row spacing of bolts (cables) is 700×700 mm. The direct roof of the 825 fully-mechanized caving face is mudstone with an average thickness of 1.84 m; the basic roof is siltstone with an average thickness of 7.00 m.

[0065] As Figure 1 shown, the surrounding rock stability control method for the end coal caving area based on numerical simulation includes the following steps:

[0066] S1. Analysis of the surrounding rock bearing structure of the roadway in the end coal caving area

[0067] Before the start of coal caving in the broken coal and rock mass in the end area of the fully-mechanized caving face, due to the lack of movement space, the coal and rock mass always keep closely biting with each other. However, with the progress of coal caving, the originally closely biting coal and rock mass around the roadway gradually loosens as the top coal is discharged; when it loosens to a certain extent, the bearing structure formed by the close biting of the coal and rock mass around the roadway will become unstable, making the difficulty of passive support in the roadway increase. If the passive support capacity in the roadway is insufficient, the roadway has the risk of instability, leading to serious safety accidents. Therefore, it is very necessary to ensure the stability of the end surrounding rock. When carrying out end coal caving, the coal and rock mass in the end area of the working face has been broken into loose blocks and can be approximated as a loose stratum. Combining with the Prandtl equilibrium arch theory, at this time, due to the redistribution of the surrounding rock stress, a parabolic pressure arch, that is, an equilibrium arch, will be formed above the roadway; and along the direction with an angle of with the horizontal direction, 2 sliding surfaces will be formed; the range formed by the equilibrium arch and the sliding surfaces is essentially the range where the surrounding rock may be damaged, as Figure 2 shown.

[0068] The suspension effect of the cable is to anchor the easily caving rock mass in the natural equilibrium arch in the deep rock. The height of the natural equilibrium arch is expressed as:

[0069]

[0070] where, σ v is the vertical pressure at the top of the roadway, MPa; γ is the unit weight of the surrounding rock, kN / m3 ; h is the height of the equilibrium arch, m; B is the semi-span of the equilibrium arch, m; b is the semi-span of the roadway, m; h 0 is the height of the roadway, m; is the calculated friction angle of the surrounding rock, °; f is the Proctor's surrounding rock strength coefficient, as follows:

[0071]

[0072] Among them, R c is the uniaxial saturated compressive strength of the surrounding rock, MPa; c is the cohesion of the surrounding rock, MPa;

[0073] From Eqs. (1) and (2), the height of the natural equilibrium arch h = 8.08 m can be obtained. That is, at this time, the entire bolt (cable) is inside the natural equilibrium arch, and the coal and rock mass inside it will be released, and the suspension effect of the cable will fail. However, before the loose coal and rock mass in the natural equilibrium arch is released, the rock masses around the end of the bolt (cable) still bite tightly with each other, and the active support effect will not fail. Moreover, under the reinforcement effect of the pre-tightening force of the bolt (cable), the broken coal and rock mass in the equilibrium arch will form a stable load-bearing structure within the direct action range of the bolt support, that is, the load-bearing arch. The essence of the formation of the load-bearing arch is due to the active support effect generated by the pre-tightening force applied by the bolt, combined with the passive support to expand the support range, thereby improving the stability of the inlaid composite force chain structure between the fractured rock masses.

[0074] Based on the Mohr-Coulomb criterion, the frictional shear force of the fractured rock mass is expressed as:

[0075]

[0076] Among them, σ n is the normal stress between the rock masses, MPa.

[0077] The load-bearing arch has a load-bearing effect due to the biting force F i formed by the mutual extrusion of the crushed stones, which is sufficient to balance the self-weight G of the crushed stones. As Figure 3 shown, the calculation formula is:

[0078] F i = F l μ (4)

[0079] Among them, F l is the lateral restraint force on the crushed stones at the arch crown, N; μ is the equivalent friction factor between the crushed stones at the arch crown;

[0080] As the top coal is continuously released, the coal and rock mass at the end of the anchor rod (cable) will loosen first, causing the active support of the anchor rod (cable) to fail. However, under the dual effects of passive support in the tunnel and compaction of the broken coal and rock mass above, the gravel at the foot of the bearing arch has not loosened and is still in a tightly biting state, with bearing capacity. Only when the gravel at the foot of the bearing arch loosens and the biting force between the gravel is unable to bear its own weight, the bearing arch will be damaged and the tunnel will lose its bearing capacity. Therefore, in order to ensure the safety of the tunnel when releasing coal at the end, the stability of the bearing arch must be ensured first.

[0081] The range of bearing arch formation is within the direct action range of anchor support, which is:

[0082] h p =αβL (5)

[0083] Among them, h p is the height of the bearing arch, m; α is the correction coefficient of the bearing arch height, and 0<α<1, the value is 2 / 3; β is the safety factor, 1≤β≤1 / α, the value is 1.1; L is the length of the direct action range of the anchor, m, the anchoring area in the tunnel is often small, usually 1.8~2.2m, and some can reach 2.5m. Remove the length of the resin roll section and take L=1.9m. Substituting into formula (5) we can get h p =1.40m.

[0084] S2. Establishment of numerical simulation model for end coal discharge

[0085] S21. Model building process

[0086] The PFC software was used to establish the numerical model for the end coal caving experiment of the gently inclined working face with different inclination angles. Since the top coal and the direct top of the end area have been broken into loose blocks during the caving stage, the top coal and the direct top are set as loose media in the ideal state, with cohesion of 0, initial velocity and displacement of 0. When caving coal at the end, the upper and lower ends of the working face are far apart, and the two will not affect each other. Therefore, the models can be established for the upper and lower end areas of the working face respectively. The following two basic principles should be followed when considering caving coal at the end: ① The middle frame of the working face is caving coal first, and after the top coal in the upper space of the middle frame is completely caving, coal is caving from the middle of the working face to both ends; ② Close the window when seeing gangue. The lower end of the model is arranged with 3 transition frames and 7 intermediate frames, numbered 1 to 10; the upper end is arranged with 3 transition frames and 17 intermediate frames, numbered 89 to 108. Among them, frames 1 and 3 are transition frames at the lower end of the working surface, and frames 106 and 108 are transition frames at the upper end of the working surface.

[0087] The roadway support combines active support and passive support. Considering that the actual generation of bolts (cables) in two-dimensional simulation will produce a limited small boundary, which has an obstructive effect on the flow of top coal, therefore, according to the actual bolt spacing, a pre-tightening force is applied using the geometry module for equivalent substitution, and the pre-tightening force is 100 KN. The bolts (cables) in the roadway are numbered as the middle cable mc, and the right half of the roadway is numbered clockwise starting from the middle cable as r 1 ~r 7 , and the left half of the roadway is numbered counterclockwise starting from the middle cable as l 1 ~l 7 ; The passive support uses Fish language to generate wall elements to simulate the shed, as Figure 4 shown. The entire coal caving process is detected in real time. When the rock mass at the end of the bolt (cable) is damaged, the active support fails, and the process is as Figure 5 shown.

[0088] The initial state of the model includes six parts: the 82 coal seam, the parting layer, the 81 coal seam, the immediate roof, the main roof, and the overlying rock. The total thickness of the parting layer and the 81 coal seam is 4.78 m, the thickness of the immediate roof is 1.84 m, and the thickness of the main roof is 7 m. To correct the particle size difference between the simulated circular particles and the actual irregular top coal blocks and simplify the number of circular particles, the particle size is selected referring to the field coal and rock diameter as shown in Table 1. The basic physical parameters of the particles are set according to the physical parameters of the coal and rock mass in the working face, and the movement behavior of the particles under the selected mesoscopic parameters is verified to be consistent with the macroscopic conditions through the natural angle of repose calibration experiment, as shown in Table 2. Among them, the density of the overlying rock is converted using Fish language according to Equation (6) to ensure that the equivalent load applied to the overlying rock layer remains unchanged during the experiment. The equivalent load is determined based on the key stratum theory, and the value is 367.72 KPa:

[0089]

[0090] Among them, P is the load of the overlying rock layer, kPa; g is the acceleration of gravity, m / s 2 ; h is the height in the rectangular area of the overlying rock in the two-dimensional model, m.

[0091] Table 1 Particle radius

[0092] Rock stratum name Coal pillar / m End area / m Middle part of the working face / m Overlying strata 0.25～0.40 0.25～0.40 0.25～0.40 Main roof 0.25～0.40 0.25～0.40 0.25～0.40 Immediate roof 0.30～0.35 0.25～0.30 0.15～0.20 81 Coal seam 0.20～0.25 0.15～0.20 0.10～0.15 Intercalated gangue layer 0.16～0.20 0.14～0.18 0.12～0.16 82 Coal seam 0.20～0.25 0.15～0.20 0.10～0.15

[0093] Table 2 Physical and mechanical parameters of the simulated coal and rock mass

[0094] Rock stratum name <![CDATA[Density / (kg·m -3 )]]> <![CDATA[Normal stiffness / (GN·m -1 )]]> <![CDATA[Shear stiffness / (GN·m -1 )]]> Coefficient of friction Overlying rock stratum ρ 0.4 0.4 0.4 Main roof 2550 0.4 0.4 0.4 Immediate roof 2030 0.3 0.3 0.3 81 Coal seam 1500 0.2 0.2 0.1 Intercalated gangue layer 2030 0.3 0.3 0.3 82 Coal seam 1500 0.2 0.2 0.1

[0095] S22. Experimental scheme

[0096] On the basis of following the two basic principles of end coal placing, the end coal placing experiment was designed as follows: for working faces with inclination angles of 8°, 17° and 25° respectively, coal was placed in the middle first, and then at the upper and lower ends.

[0097] S3. Analysis of simulation results

[0098] 1. Temporal and spatial analysis of active support

[0099] In order to analyze the temporal and spatial differences of active support in the tunnel, during the coal caving process, a 5×10 4 The time step is a basic time unit, and the effective period of active support during the coal caving process at the upper and lower ends of the working face with different inclination angles is counted respectively. The average number of coal caving steps for the No. 4 intermediate frame is 1×10 6 During the initial coal caving process from the beginning to the end of the No. 4 intermediate frame, the active support function in the tunnel began to fail one after another.

[0100] (1) Lower end

[0101] The temporal and spatial effect statistics of the active support at the lower end are as follows: Figure 6 As shown. 2 Anchor cable and 3 The time nodes of failure of active anchor support are generally not much different, and all fail at the beginning of coal caving, with the number of operation steps being about 1×10 5 Time step. 2 Anchor cable and 3 After the bolt fails, the middle anchor cable mc begins to fail, and as the inclination of the working face increases, the time required for mc to fail becomes longer. In addition, the greater the inclination of the working face, the greater the horizontal component of force on the lower end, making the r 1 Anchor failure.

[0102] Therefore, although there are differences in the time nodes of failure of anchor rods (cables) in the roadway during the initial coal caving process at the lower end of the working face with different inclination angles, the active support of anchor rods (cables) in the middle and right half of the roadway basically fails after the coal caving is completed. In order to ensure that the bearing arch around the roadway can still maintain its bearing capacity after the active support fails, the passive support in the middle and right half of the roadway should be strengthened before coal caving, ensuring that the advanced support in the roadway is strictly connected to the top and has sufficient strength, ensuring the close engagement of the coal and rock mass at the arch foot of the bearing arch, and preventing the failure of the bearing arch from causing the roadway to become unstable.

[0103] (2) Upper end

[0104] The statistics of the temporal and spatial effects of active support at the end of the working face are as follows: Figure 7 As shown. The anchor cables mc and l in the middle of the upper end tunnel 3The spatio-temporal characteristics of the failure of the active support of the bolt are not significantly affected by the dip angle of the working face, and all fail after the coal is discharged in the middle of the working face to the 101st support. l 1 The bolt and l 2 The spatio-temporal characteristics of the failure of the active support of the cable bolt are most affected by the dip angle of the working face; among them, l 2 The time node of the cable bolt failure is earlier than l 1 the bolt, and the greater the dip angle of the working face, l 2 the farther the cable bolt is affected by the coal discharge in the middle of the working face. In the roadway of the working face with an 8° dip angle, the l 1 bolt does not fail. In the roadway of the working face with a 17° dip angle, the l 1 bolt fails after the coal is discharged in the middle of the working face to the 103rd support. In the roadway of the 25° working face dip angle, the l 1 bolt fails after the coal is discharged in the middle of the working face to the 105th support. This shows that as the dip angle of the working face increases, the influence distance of the coal discharge in the middle of the working face on the l 1 bolt in the roadway decreases, l 1 and the affected range of the bolt is negatively correlated with the dip angle of the working face. Similarly, l 2 the affected range of the cable bolt is positively correlated with the dip angle of the working face.

[0105] Therefore, as the coal discharge in the middle of the working face progresses, the active support of the bolts (cable bolts) in the middle and the left half of the upper-end roadway will fail successively before the coal is discharged to the transition support position. The distance of the bolts (cable bolts) affected and the sequence of the bolt (cable bolt) failure time are both related to the dip angle of the working face. When the coal is discharged to the transition support position, the active support of the bolts (cable bolts) in the middle and the left half of the roadway basically fails. Similar to the situation of the lower end, to ensure the safety and stability of the roadway, before the coal is discharged at the upper end of the working face, it is necessary to strengthen the passive support in the roadway to ensure the stability of the arch feet of the load-bearing arch. It should be noted that when the coal is discharged in the middle of the working face near the upper end of the working face, starting from the position 10 - 13 center distances of the supports ahead, the upper end of the working face begins to be continuously affected by the coal discharge in the middle of the working face. Therefore, when approaching the upper end of the working face and during subsequent coal discharge operations, the passive support in the middle and the left half of the roadway can be strengthened in advance to ensure the stability of the load-bearing arch.

[0106] 2. Analysis of the flow pattern of coal and gangue in the coal discharge area at the ends

[0107] Affected by the coal seam dip angle, there are differences in the flow laws of the coal and gangue layers at the upper and lower ends of the working face. Therefore, the coal discharge processes at the upper and lower ends are analyzed separately. When discharging coal at the ends, taking 1×10 4 time steps as 1 basic time unit to observe and evaluate the flow and migration characteristics of the coal and rock mass and the evolution law of the surrounding rock structure of the roadway during the coal discharge process at the ends.

[0108] (1) Lower end

[0109] Coal is discharged from the lower end of the working face with different dip angles, and the displacement field results are as follows Figure 8 As shown. During the coal discharge process at the lower end of the working face, the coal discharge from the No. 3 support is generally safe, and with the increase of the dip angle of the working face, the safety increases; when discharging coal from the No. 2 support, it is very easy to cause gangue to flow through, which is likely to lead to misjudgment of closing the window of the support and resulting in loss of top coal; after the coal discharge from the No. 1 support, the load-bearing arch becomes loose and there is a risk of roadway instability. It should be noted that after the coal discharge from the No. 3 support of the working face with an 8° dip angle, there is a slight loosening at the edge of the load-bearing arch, which is significantly different from the state of the load-bearing arch after the coal discharge from the No. 3 support of the other two dip angle working faces. Analyzing the coal discharge process of the No. 3 support of the 8° dip angle working face according to the time node, the coal and rock mass within the load-bearing arch has already displaced after the coal discharge from the middle support of the working face at the lower end of the 8° dip angle. During the coal discharge process, the area where displacement occurs within the load-bearing arch does not continue to develop and the shape remains consistent all the time, and the load-bearing arch can still remain stable. For safety reasons, the support in the roadway can be strengthened before coal discharge to prevent the instability of the load-bearing arch.

[0110] Therefore, when discharging coal at the same position on working faces with different dip angles, the smaller the dip angle of the working face, the closer the coal and rock mass that generates displacement is to the edge of the load-bearing arch, that is, the failure speed of the load-bearing arch at the lower end of the working face is negatively correlated with the dip angle of the working face. In addition, after the coal discharge from the No. 1 support, compared with the 8° and 17° dip angle working faces, in the 25° dip angle working face, a tree-branch-shaped sudden intrusion displacement penetrates through the load-bearing arch within the load-bearing arch range, causing the coal and rock mass at the position slightly to the left of the middle to become loose. It is speculated that it may be because the lower end of the working face with a larger dip angle is more dense, and the coal and rock mass bite more tightly after fragmentation, so it is easier to form a local load-bearing structure; after the coal body above the coal discharge port is discharged, a free face is generated on the side of the originally dense coal and rock mass, and the interaction force between the coal and rock mass causes the coal and rock mass to move towards the free face and generate displacement, resulting in the destruction of the local load-bearing structure, causing the coal and rock mass that was originally stable under the action of the local load-bearing structure to move, and finally triggering the sudden displacement of the coal and rock mass within the load-bearing arch area. Therefore, when discharging coal at the lower end of the working face with a larger dip angle, it is also necessary to appropriately strengthen the support in the roadway.

[0111] (2) Upper end

[0112] Coal is discharged from the upper end of the working face with different dip angles, and the coal discharge displacement field results are as follows Figure 9As shown. After the middle coal caving is completed on the working faces with different dip angles, the load-bearing arches have all loosened to varying degrees. In terms of the loosening range of the load-bearing arch and the displacement of the coal and rock mass within the arch, the 25° dip angle working face > the 17° dip angle working face > the 8° dip angle working face. According to the process of middle support coal caving, when the middle support coal caving on the working faces with different dip angles gradually approaches the end area, a fork-shaped displacement will appear in advance within the space between the upper part of the left half of the load-bearing arch and the coal caving opening. Different from the sudden intrusion displacement at the lower end, the fork-shaped displacement at the upper end does not penetrate the load-bearing arch. As the coal caving progresses, the coal caving funnel continues to develop, and the coal and rock mass between the load-bearing arch and the coal caving opening gradually generates displacement and approaches the fork until it finally penetrates. Comparing the time nodes when the fork-shaped displacement appears on the working faces with different dip angles, the 25° dip angle working face is the earliest, the 17° working face is the second, and the 8° dip angle working face is the latest; comparing the penetration speed of the coal and rock mass between the load-bearing arch and the coal caving opening after displacement, the 8° dip angle working face is the fastest, the upper end of the 17° working face is the second, and the upper end of the 25° dip angle working face is the slowest.

[0113] Therefore, the middle coal caving on the working face has an advanced influence on the loosening of the upper end load-bearing arch, and the advanced influence distance is 6 - 11 center distances of the supports. As the dip angle of the working face increases, the advanced influence distance will also become farther, that is, the advanced influence distance is positively correlated with the dip angle of the working face. The reason is related to the stronger movement trend of the coal and rock mass at the upper end of the working face with a larger dip angle. In addition, as the coal caving progresses, the continuous development of the coal caving funnel above the coal caving opening will cause the range of the coal and rock mass that generates displacement between the load-bearing arch and the coal caving opening to continuously expand. However, due to the different lateral restraint effects of the working face floor, there are differences in the penetration speed of the coal and rock mass, and the penetration speed of the 8° dip angle working face is the fastest, and that of the 25° dip angle working face is the slowest. Compared with the lower end of the working face, the load-bearing arches at the upper ends of the working faces with different dip angles have loosened to varying degrees before the coal caving of the transition support, indicating that there is a risk of instability in the upper end roadway after the middle support coal caving on the working face is completed. Therefore, before the coal caving on the working face approaches the upper end, the support in the roadway should be strengthened. In addition, due to the phenomenon of "advanced penetration displacement" occurring at the upper end of the working face during the coal caving process, it is necessary to be vigilant against the sudden emptying of the upper end of the working face during the coal caving operation.

[0114] 3. Analysis of the evolution of the bearing capacity of the surrounding rock bearing structure of the roadway

[0115] Although the above content analyzes the displacement field during the coal caving process on the working face, the displacement field only reflects the dynamic loosening process of the load-bearing arch and is not sufficient to show the dynamic change process of the bearing capacity of the load-bearing arch. Therefore, it is necessary to analyze the force chain field during the coal caving process. Force chains are the main form of transmitting loads in discontinuous and granular media and are also the bridge connecting the macroscopic mechanical behavior and mesoscopic action mechanism of granular media. Using the existing force chain classification method, force chains are classified into three grades according to strength: strong force chains, secondary strong force chains, and weak force chains.

[0116]

[0117] Among them, F is the contact force between particles, in N; is the average contact force, in N.

[0118] The combined force chain structure formed by the three types of strength force chains breaks and regenerates continuously throughout the coal caving process. Generally, it shows that after the strong force chain that plays a major supporting role breaks, the coal body becomes unstable, and then during the process of re-stabilization, the strong force chain regenerates again, accompanied by the secondary strong force chain and the weak force chain filling the strong force chain framework to assist in bearing. The specific process is as follows: ① The strong force chain is connected, the biting state between the coal and rock masses is good, an effective transmission channel for the contact force between the coal and rock masses can be formed, the bearing capacity of the bearing arch is good, and the continuous stability of the roadway can be ensured; ② The strong force chain breaks, the biting between the coal and rock masses becomes loose, the transmission channel for the contact force between the coal and rock masses gradually disappears, the bearing capacity of the bearing arch weakens, and there is a risk of roadway instability; ③ The strong force chain is reconstructed, the biting between the coal and rock masses no longer exists, the transmission channel for the contact force between the coal and rock masses completely disappears, the bearing arch loses its bearing capacity, the roadway becomes unstable, and the coal and rock masses come into contact again under the action of gravity and tend to self-stabilize and reconstruct the strong force chain. In the numerical model, the Fish language is used to screen and distinguish different strength force chains, and the change process of the force chain field during the end coal caving process is analyzed.

[0119] (1) Lower end

[0120] The change of the force chain field during the coal caving process at the lower end of the working face is studied, and the results are as Figure 10 shown. After the coal caving of the No. 3 support is completed, the strong force chains in the combined force chain structure of the bearing arch of the working face with three dip angles are all in the state of strong force chain connection; during the coal caving process of the No. 2 support, due to the intrusion of gangue, the window of the support is closed in advance. Although the overall shape of the combined force chain structure has basically not changed, after the coal caving is completed, the strong force chains in the combined force chain structure of the bearing arch of the working face with three dip angles are all in the state of strong force chain breakage; after the coal caving of the No. 1 support is completed, the strong force chains in the combined force chain structure of the bearing arch of the working face with three dip angles are all in the state of strong force chain reconstruction. Among them, after the coal caving of the No. 3 support is completed, the good degree of the connection state of the strong force chains in the combined force chain structure of the bearing arch is in the order of the working face with a 25° dip angle, the working face with a 17° dip angle, and the working face with an 8° dip angle.

[0121] Therefore, during the coal caving process at the lower end of the working face, after the No. 3 frame is finished caving, the coal and rock masses in the bearing arch are tightly engaged, the bearing capacity of the bearing arch is good, and the roadway stability is good; after the No. 2 frame is finished caving, the coal and rock masses in the bearing arch are loosened, the bearing capacity of the bearing arch is weakened, and there is a risk of instability in the roadway; after the No. 1 frame is finished caving, the bearing arch structure is unstable, loses its bearing capacity, and the roadway is unstable. It is worth noting that the bearing capacity of the bearing arch is closely related to the tightness of the coal and rock mass engagement in the arch, and the degree of engagement of the coal and rock mass at the lower end is affected by the inclination of the working face. The larger the inclination of the working face, the tighter the engagement between the coal and rock masses, the stronger the bearing capacity of the bearing arch, and the higher the safety of the roadway. In other words, the bearing capacity of the bearing arch at the lower end of the working face is positively correlated with the inclination of the working face. At the same time, combined with the above displacement field analysis results, the coal caving danger level of the lower end of the working face from the No. 3 frame to the No. 1 frame can be divided into the following zones: ① No. 3 frame, coal caving safety zone; ② No. 2 frame, coal caving risk zone; ③ No. 1 frame, coal caving danger zone ( Figure 11 ).

[0122] (2) Upper end

[0123] The changes of force chain field during coal caving at the end of working faces with different inclination angles were studied. However, after the coal caving of the No. 106 frame was completed, the combined force chain structure of the bearing arch at the end of the working faces with different inclination angles was already in a state of strong chain fracture. In order to clarify the dynamic change process of the bearing capacity of the bearing arch during the coal caving of the No. 106 frame, the combined force chain state of the bearing arch corresponding to different time nodes was counted. 5 At the 17° inclination working face, the total coal caving of the 106th frame was 2.1×10 5 At the 25° inclination working face, the total coal caving of the 106th frame was 2.4×10 5 The combined force chain state of the bearing arch corresponding to a specific time node is as follows: Figure 12 As shown. The 8° inclined working face is running at 1×10 4 After the time step, the combined force chain structure of the bearing arch is in a state of force chain fracture; the 17° inclined working face is running at 1.9×10 5 At this time, when the coal caving progress was at 95%, the combined force chain structure of the bearing arch was in a state of force chain fracture; the 25° inclined working face was caving coal at the 106th frame to 1.5×10 5 After the time step, when the coal caving progress is at 65%, the combined force chain structure of the bearing arch is in a state of strong force chain fracture. According to the order of occurrence of the strong force chain fracture state of the combined force chain structure in the bearing arch, it is 25° inclined working face, 17° inclined working face, and 8° inclined working face. At the same time, when the combined force chain structure of the bearing arch is in a state of strong force chain fracture, the intensity of the strong force chain fracture is from strong to weak, which is 25° inclined working face, 17° inclined working face, and 8° inclined working face.

[0124] Therefore, during the top coal caving process at the upper end of the working face, the dynamic change process of the combined force chain structure of the load-bearing arch is similar to that at the lower end of the working face, and it will go through three stages: strong force chain connection - strong force chain fracture - strong force chain reconstruction. The difference is that the time node and spatial position of the combined force chain structure of the load-bearing arch at the upper end of the working face reaching the strong force chain fracture state are earlier than those at the lower end. During the coal caving period of the 106th support, the combined force chain structure will reach the strong force chain fracture state, that is, the complete caving of the top coal in the area above the 106th support will cause a significant reduction in the load-bearing capacity of the load-bearing arch, and there is a risk of roadway instability. Moreover, the greater the dip angle of the working face, the earlier the time node for the combined force chain structure to reach the strong force chain fracture state, and the more severe the fracture degree of the strong force chain. This is contrary to the reason why the load-bearing capacity of the load-bearing arch at the lower end of the working face increases with the increase of the dip angle of the working face. The tightness of the bite between the coal and rock masses at the upper end of the working face will decrease with the increase of the dip angle of the working face, and the load-bearing capacity of the load-bearing arch will also decrease with the increase of the dip angle of the working face. In other words, the load-bearing capacity of the load-bearing arch at the upper end of the working face is negatively correlated with the dip angle of the working face. Combining with the above displacement field analysis results, it is concluded that for safety, top coal caving at the upper end of the working face should only be carried out at the 106th support and cannot be completely caved. The coal caving section within the support can be divided into a safe coal caving section and a risky coal caving section according to the degree of danger, as Figure 13 shown.

[0125] (3) Determination of the safe coal caving height within the support

[0126] To determine the specific height of the safe coal caving section, the ultimate coal caving progress of the 106th support on working faces with dip angles of 8°, 12°, 17°, 21° and 25° was statistically analyzed respectively, and the ultimate coal caving progress equation was fitted, as Figure 14 shown. The variation law of the ultimate coal caving progress of the 106th support with the dip angle is consistent with the exponential function. Therefore, when the dip angle of the working face increases to a certain extent, the ultimate coal caving progress of the 106th support will decrease exponentially, that is, when the dip angle of the working face is too large, the top coal at the end should not be caved to ensure operation safety.

[0127] Based on this, the height of the safe coal caving section during the coal caving process of the 106th support is deduced as follows:

[0128] h s =[0.97901 - 0.0158exp(0.13103α)]h d (8)

[0129] where, h s is the height of the safe coal caving section, m; h d is the coal caving height, m; α is the dip angle of the working face, °.

[0130] As Figures 15 - 17 shown, a device for a surrounding rock stability control method in the end coal caving area based on numerical simulation includes an end coal caving support;

[0131] According to the above simulation results, taking the 825 face of a coal mine as an example and substituting the data for calculation, we get: the limit caving progress of the 106th support at the upper end of the 825 face is 92%, the height of the safe caving section is 4.40 m. After removing the thickness of the upper parting layer, the remaining caving height of the top coal is 2.46 m. By comprehensively comparing the top coal recovery amounts with and without caving at the face ends of the 825 working face, after recovering the top coal in the face end area, the top coal recovery amount is increased by 1.8% compared with only caving in the middle of the working face. In addition, when conducting sectional caving within the support at the upper end of the working face, in order to accurately divide the height of the safe caving section and control the end time of caving, a top coal displacement tracker can be arranged at the interface between the safe caving section and the caving risk section above the 106th support. By real-time tracking the displacement trajectory of the top coal, the sectional caving within the support at the upper end of the working face can be achieved quickly, accurately, safely and efficiently.

[0132] At present, for caving at the face ends, there is a problem that there are many mechanical equipment in the face end area, but the existing caving supports do not have enough space to conduct caving in the upper and lower face end areas. Therefore, considering the time and cost of equipment research and development, the existing hydraulic supports can be directly modified to meet the caving requirements at the face ends. Specific solutions: ① Set windows on the shield beam 1, and configure hydraulic cylinders 4 on both sides. The window part is replaced with a high-strength flexible metal mesh 2 instead of the traditional structure; ② Control the extension and retraction of the metal mesh by the telescopic operation of the hydraulic cylinder 4 to achieve precise control of the coal discharge opening 8 and complete the caving operation at the face ends; ③ Integrate the control system of the hydraulic cylinder 4 into the intelligent electro-hydraulic control module to realize the intelligent comprehensive caving of the face end hydraulic supports and improve the efficiency and accuracy of mining operations. The expected size of the window on the shield beam 1 is 1200 mm in width × 1500 mm in height. The two hydraulic cylinders 4 cooperate to push the high-strength flexible metal mesh 2 to control the opening and closing of the coal discharge opening 8. In addition, a mesh winding device 6 is also equipped below the shield beam 1, and a coal breaking knife 3 is used to break coal blocks at the end of the metal mesh. A coal breaking knife groove 5 adapted to the coal breaking knife 3 is arranged on the shield beam 1, a mesh winding device 6 is arranged below the shield beam 1, a mesh winding guide plate 7 is arranged on one side of the mesh winding device 6, and a coal discharge opening 8 is arranged on the side of the flexible metal mesh 2 close to the coal breaking knife 3.

[0133] Through the verification and analysis of the above embodiments, the following conclusions are obtained:

[0134] (1) Before caving starts in the face end area, the active support effects of the upper and lower face end roadways of the working face are basically invalid under the influence of caving in the middle of the working face, and there are differences in the time nodes and spatial positions of invalidation. The active support effect of the upper face end roadway will be affected in advance by caving in the middle of the working face, and the advance influence distance is 10 - 13 center distances of the supports, and the advance influence distance is positively correlated with the working dip angle.

[0135] (2) During the coal caving process in the end region, the load-bearing arch at the lower end does not loosen after the coal caving in the middle of the working face is completed. It gradually loosens with the coal caving of the transition support, and when the dip angle of the working face is large, sudden intrusion displacement occurs to damage the load-bearing arch. The load-bearing arch at the upper end loosens in advance under the influence of the middle coal caving. The advance influence range is 6 - 11 center distances of the supports, and the coal and rock mass within the advance influence range gradually loosens and penetrates. During the coal caving process, the flow characteristics of the coal and rock mass need to be fully considered, and appropriate protective measures should be taken.

[0136] (3) The coal caving schemes at the working face ends are the overall zoning at the lower end and the in-frame segmentation at the upper end. During the coal caving process at the lower end, it is overall divided into a coal caving safety area, a coal caving risk area, and a coal caving danger area. During the coal caving process at the upper end, the top coal above the transition support is divided into a coal caving safety section and a coal caving risk section. The height of the coal caving safety section decreases exponentially with the increase of the dip angle of the working face.

[0137] (4) By transforming the existing hydraulic supports, not only can the cost be saved, but also precise coal caving control can be achieved in cooperation with the top coal tracker. The successful caving of the end top coal increases the top coal recovery rate by 1.8%.

[0138] Therefore, the present invention adopts the above-mentioned method and device for controlling the stability of surrounding rock in the end coal caving area based on numerical simulation to study the coal and rock flow law, the stability of end surrounding rock, and the reasonable end coal caving range during the coal and rock caving process, which not only improves the safety and efficiency of coal mine mining, but also reduces the cost and promotes technological innovation.

[0139] Finally, it should be noted that the above embodiments are only used to illustrate the technical solutions of the present invention and are not intended to limit them. Although the present invention has been described in detail with reference to the preferred embodiments, those of ordinary skill in the art should understand that they can still modify or equivalently replace the technical solutions of the present invention, and these modifications or equivalent replacements cannot make the modified technical solutions deviate from the spirit and scope of the technical solutions of the present invention.

Claims

1. A method for controlling the stability of surrounding rock in the end caving area based on numerical simulation, characterized in that: The following steps are involved: S1. Analyze the bearing structure of the surrounding rock of the end coal caving area; S2. Use PFC software to establish a numerical simulation model for end coal discharge; S3. Analyze the simulation results of the end coal caving numerical simulation model based on S2, including the spatiotemporal analysis of active support, the flow morphology analysis of coal gangue in the end coal caving area, and the evolution analysis of the bearing capacity of the tunnel surrounding rock bearing structure.

2. The method for controlling the stability of surrounding rock in the end caving area based on numerical simulation according to claim 1 is characterized in that: Step S1 is specifically as follows: Before coal caving, the coal and rock bodies in the end area of ​​the fully mechanized caving face are tightly interlocked with each other. As the coal caving progresses, the coal and rock bodies around the roadway, which were originally tightly interlocked, gradually loosen with the release of the top coal. When the end coal caving is carried out, the coal and rock bodies in the end area of ​​the working face have been broken into loose blocks. Combined with the Proctor equilibrium theory, at this time, due to the redistribution of the surrounding rock stress, a parabolic pressure arch will be formed above the roadway, which is a balanced arch; and the angle between the horizontal direction and the pressure arch is In the direction, two sliding surfaces will be formed; the range formed by the balance arch and the sliding surface is the range where the surrounding rock is destroyed; The height of the natural equilibrium arch is expressed as: Among them, σ v is the vertical pressure at the top of the tunnel; γ is the bulk density of the surrounding rock; h is the height of the balance arch; B is the half span of the balance arch; b is the half span of the tunnel; h0 is the height of the tunnel; is the friction angle calculated for the surrounding rock; f is the solidity coefficient of the surrounding rock, as follows: Among them, R c is the uniaxial saturated compressive strength of surrounding rock; c is the cohesion of surrounding rock; Before the loose coal and rock mass in the natural balance arch is released, the rock mass around the end of the anchor rod or anchor cable is still tightly engaged with each other, and the active support function will not fail. Moreover, the broken coal and rock mass in the balance arch will form a stable bearing structure within the direct action range of the anchor rod support under the reinforcement of the pre-tightening force of the anchor rod or anchor cable, which is a bearing arch. Based on the Mohr-Coulomb criterion, the friction shear force of the broken rock mass is expressed as: Among them, σ n is the normal stress between rock masses. The bearing arch has a bearing function due to the bite force F formed by the crushed stones squeezing each other. i Sufficient to balance the gravel's own weight G, the calculation formula is: F i =F l m (4) Among them, F l is the lateral restraint force on the gravel of the vault; μ is the equivalent friction coefficient between the gravel of the vault; Therefore, when the end coal is placed, the safety of the roadway is ensured by the stability of the bearing arch. The range of the bearing arch is within the direct action range of the anchor support, which is: h p =αβL (5) Among them, h p is the bearing arch height; α is the bearing arch height correction coefficient, and 0<α<1; β is the safety factor, 1≤β≤1 / α; L is the length of the direct action range of the anchor rod.

3. The method for controlling the stability of surrounding rock in the end caving area based on numerical simulation according to claim 1 is characterized in that: When coal is placed at the end in S2, the distance between the upper end and the lower end of the working face is far, and the two will not affect each other. Models are established for the upper end and lower end areas of the working face respectively. The tunnel support adopts a combination of active support and passive support. Considering that the actual generation of anchor rods or anchor cables in the two-dimensional simulation will produce a limited small boundary, the geometry module is used to apply pre-tightening force according to the actual spacing between anchor rods for equivalent replacement. The anchor rods or anchor cables in the tunnel are numbered as the middle anchor cable mc, and the right part of the tunnel is numbered r1~r7 clockwise from the middle anchor cable, and the left half of the tunnel is numbered l1~l7 counterclockwise from the middle anchor cable. The passive support uses the Fish language to generate wall units to simulate the shed.

4. The method for controlling the stability of surrounding rock in the end caving area based on numerical simulation according to claim 1 is characterized in that: The initial state of the model in S2 includes six parts: 82 coal seams, interbedded gangue layers, 81 coal seams, direct roof, basic roof and overburden. The overburden density is converted using the Fish language according to the following formula: Where P is the overburden load; g is the gravitational acceleration; and h is the height of the rectangular area of ​​the overburden in the two-dimensional model.

5. The method for controlling the stability of surrounding rock in the end caving area based on numerical simulation according to claim 1 is characterized in that: The spatiotemporal analysis of active support in S3 is to analyze the spatiotemporal difference characteristics of active support in the roadway. In the process of coal caving, the effective period of active support in the upper and lower ends of the working face with different inclination angles is counted as a basic time unit, including: Lower end: During the initial coal caving process at the lower end of working faces with different inclination angles, although there are differences in the time nodes at which the anchor rods or anchor cables in the roadway fail, after the coal caving is completed, the active support of the anchor rods or anchor cables in the middle and right parts of the roadway basically fails. Before coal caving, the passive support in the middle and right parts of the roadway is strengthened to ensure that the bearing arch around the roadway can still maintain its bearing capacity after the active support fails. Upper end: When placing coal to the transition frame position, the active support functions of the anchor rods or cables in the middle and left parts of the tunnel are basically ineffective. Before placing coal at the upper end of the working face, it is necessary to strengthen the passive support in the tunnel to ensure the stability of the arch foot.

6. The method for controlling the stability of surrounding rock in the end caving area based on numerical simulation according to claim 1 is characterized in that: The flow pattern analysis of coal gangue in the end caving area of ​​S3 is to observe and evaluate the flow and migration characteristics of coal and rock mass and the evolution law of tunnel surrounding rock structure during the end caving process, with a basic time unit, including: Lower end: When coal is placed at the same position on working faces with different inclination angles, the failure rate of the bearing arch at the lower end of the working face is negatively correlated with the inclination angle of the working face; Upper end head: Coal caving in the middle of the working face has a leading effect on the loosening of the upper end bearing arch, and the leading influence distance is positively correlated with the inclination angle of the working face.

7. The method for controlling the stability of surrounding rock in the end caving area based on numerical simulation according to claim 1 is characterized in that: The evolution analysis of the bearing capacity of the surrounding rock bearing structure of the S3 tunnel needs to analyze the force chain field during the coal caving process to show the dynamic change process of the bearing capacity of the bearing arch. The force chain is the main form of load transmission by discontinuous and bulk media. Through the force chain division method, the force chain is divided into three levels according to the strength: strong force chain, sub-strong force chain and weak force chain, which are: Where F is the contact force between particles; is the average contact force; The combined force chain structure formed by the three strength force chains breaks and regenerates continuously throughout the entire coal caving process, including strong force chain connection-strong force chain breakage-strong force chain reconstruction. The Fish language is used in the model to screen and distinguish different strength force chains, and the change process of the force chain field during the end coal caving process is analyzed, including: Lower end: The bearing capacity of the bearing arch at the lower end of the working surface is positively correlated with the inclination angle of the working surface; Upper end: The bearing capacity of the upper end bearing arch on the working surface is negatively correlated with the inclination angle of the working surface.

8. A device for controlling the stability of surrounding rock in the end caving area based on numerical simulation according to any one of claims 1 to 7, characterized in that: It is used to meet the coal placing needs in the end coal placing area, including an end coal placing support, which includes a shielding beam, a window is arranged on the top of the shielding beam, hydraulic cylinders are arranged on both sides of the shielding beam, a flexible metal net is arranged in the window, and the extension and retraction of the flexible metal net are controlled by the hydraulic cylinder, a coal breaking knife is arranged at one end of the flexible metal net, a coal breaking knife groove matched with the coal breaking knife is arranged on the shielding beam, a net rolling device is arranged under the shielding beam, a net rolling guide plate is arranged on one side of the net rolling device, and a coal placing port is arranged on the side of the flexible metal net close to the coal breaking knife.

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