Method for simulating and analyzing bearing pressure distribution of fully mechanized caving face of extra-thick coal seam

Abstract

The invention relates to the technical field of mining engineering, in particular to an extra-thick coal seam fully mechanized caving face bearing pressure distribution simulation analysis method which comprises the steps of S1, bearing stability analysis; s2, crack analysis; s3, analyzing the influence of the topping operation; s4, evaluating feasibility of roof cutting roadway protection; according to the method, the stress data of the coal pillar are collected and analyzed to obtain the bearing stability of the coal pillar, the crack propagation speed is analyzed through image acquisition, so that the crack severity degree of the top plate is obtained, and the effect evaluation coefficient of the coal seam roof cutting operation is obtained according to the pressure parameter analysis of the coal seam working face after the roof cutting operation. The roadway deformation amount and the surrounding rock stability are obtained through video recording analysis, and then the feasibility of roof cutting roadway protection is obtained through analysis, so that the coal seam fully mechanized caving face bearing pressure influence evaluation index is comprehensively analyzed, targeted measures can be better taken to guarantee operation safety, the mining technology and process can be optimized, and the mining efficiency is improved.

Description

Technical Field

[0001] The present invention relates to the technical field of mining engineering, and more specifically, to a method for simulating and analyzing the abutment pressure distribution in a fully-mechanized caving face of an extra-thick coal seam. Background Art

[0002] In the field of coal mining, the fully-mechanized caving face of a coal seam is an important mining method. With the continuous deepening of coal resource mining, higher requirements are put forward for the safety and efficiency of the working face.

[0003] Understanding the abutment pressure distribution in the fully-mechanized caving face of a coal seam is crucial for ensuring mining safety and optimizing mining technology. The distribution of the abutment pressure directly affects key links such as the stability of the roadway surrounding rock, the support effect of the roof, and the advancement of the working face. Through the simulation study of the abutment pressure distribution, the variation law of the pressure under different mining conditions can be deeply analyzed, providing a scientific basis for the reasonable design of roadway layout, the selection of support methods, and the determination of mining parameters.

[0004] For example, the existing Chinese patent with the application number 201810751300.1 discloses a method for determining the spacing of pressure-relief boreholes based on the target abutment pressure. This scheme measures the original rock stress of the coal seam and the curve and peak value of the advanced abutment pressure, calculates the radius of the plastic zone of the borehole to understand the stress distribution around the borehole, then constructs pressure-relief boreholes with an increasing spacing along the mining direction of the working face, and installs borehole stress gauges in the boreholes to monitor the change of the borehole stress in real time as the working face advances, so as to directly obtain the abutment pressure at different positions of the boreholes, and can relatively accurately grasp the stress situation of the coal seam and the stress distribution around the borehole.

[0005] However, there are the following problems in the above patent: This scheme mainly focuses on the measurement and analysis of the stress distribution around the borehole. However, this is not comprehensive enough. The analysis of the coal seam requires detailed analysis and evaluation in many aspects, such as coal pillar bearing, roof cracks, roof cutting operations, etc. If a comprehensive evaluation cannot be carried out, it may lead to inaccurate grasp of the overall situation of the coal seam, and then cause a series of problems such as unreasonable mining planning, increased safety risks, and low mining efficiency, seriously affecting the smooth progress and benefits of coal seam mining. Summary of the Invention

[0006] In order to overcome the shortcomings in the background art, the embodiment of the present invention provides a method for simulating and analyzing the abutment pressure distribution in a fully-mechanized caving face of an extra-thick coal seam, which can effectively solve the problems involved in the above background art.

[0007] The object of the present invention can be achieved by the following technical solutions: A simulation analysis method for the abutment pressure distribution in a fully-mechanized caving face of extra-thick coal seams, the method comprising the following steps: S1. Bearing stability analysis: Collect stress data of coal pillars, and analyze to obtain the bearing stability of coal pillars. The stress data of coal pillars include the proportion of high-stress areas at each time point and the stress value growth rate.

[0008] S2. Crack analysis: Determine the crack start time point according to the abnormal degree of the roof ultrasonic signal at each time point, and then obtain the crack propagation speed through image acquisition and analysis, so as to obtain the severity of the roof cracks.

[0009] S3. Influence analysis of roof cutting operation: Analyze and obtain the effect evaluation coefficient of the coal seam roof cutting operation according to the pressure parameters of the coal seam working face after the roof cutting operation. The pressure parameters include the pressure uniformity degree and the change value of the pressure uniformity degree.

[0010] S4. Feasibility evaluation of roof cutting for roadway protection: When the roof collapses, analyze the roadway deformation amount and the surrounding rock stability through video recording, and then analyze the feasibility of roof cutting for roadway protection.

[0011] S5. Comprehensive analysis of abutment pressure: According to the bearing stability ξ of the coal pillar, the severity θ of the roof cracks, the effect evaluation coefficient λ of the coal seam roof cutting operation, and the feasibility χ of roof cutting for roadway protection, analyze and obtain the influence evaluation index of the abutment pressure in the fully-mechanized caving face of the coal seam and give feedback.

[0012] Preferably, the specific detection method for the stress data of the coal pillar is as follows: First step, select a number of detection points at different parts of the coal pillar according to the set spacing, and at the same time select a number of time points according to the set time interval. Collect the stress values of each detection point of the coal pillar at each time point through stress sensors, and record them as the stress values of each detection point of the coal pillar at each time point. Compare the stress values of each detection point of the coal pillar at each time point with the preset high stress value threshold, and screen out each detection point where the stress value of the coal pillar at each time point is greater than or equal to the preset high stress value threshold, and record them as each high stress value detection point of the coal pillar at each time point. Connect each high stress value detection point of the coal pillar at each time point to outline the boundary of the high stress area of the coal pillar at each time point, and obtain the area of the high stress area of the coal pillar at each time point, denoted as M i , i represents the number of the i-th time point, i = 1, 2,..., n, through the formula obtain the proportion of the high stress area of the coal pillar at each time point M 煤柱 represents the cross-sectional area of the coal pillar.

[0013] In the second step, the stress values of each detection point of the coal pillar at each time point are averaged to obtain the stress values of the coal pillar at each time point. Taking two adjacent time points as a group, denoted as each time point group, the stress values of each time point group of the coal pillar are subtracted to obtain the stress value differences of each time point group of the coal pillar. By dividing the stress value differences of each time point group of the coal pillar by the set time interval, the stress value growth rate of each time point group of the coal pillar is obtained. Furthermore, the stress value growth rate of the coal pillar is calculated through mean value calculation, denoted as V.

[0014] Preferably, the specific analysis method for the bearing stability analysis is as follows: Read the proportion of the high-stress area of the coal pillar at each time point respectively. The stress value growth rate V of the coal pillar is substituted into the formula to obtain the bearing stability ξ of the coal pillar. V' represents the preset reference value of the stress value growth rate, φ1 and φ2 respectively represent the weight factors of the proportion of the high-stress area and the stress value growth rate preset, and n represents the number of time points.

[0015] Preferably, the specific analysis method for determining the crack start time point is as follows: In the first step, the roof images at each time point are obtained through a camera. At the same time, a number of detection points are selected on the roof at a set distance, denoted as each roof detection point. A coupling agent is applied to each roof detection point on the roof, and an ultrasonic flaw detector is used to emit ultrasonic signals with a fixed intensity to each roof detection point at each time point, and a probe is used to receive the ultrasonic signals reflected from the roof. The duration from the emission to the reception of the ultrasonic signals at each roof detection point at each time point is recorded, denoted as T im , m represents the number of the mth roof detection point, m = 1, 2,..., q. At the same time, the intensity of the ultrasonic signals reflected from each roof detection point at each time point is denoted as G im , and through the formula the abnormality degree δ of the roof ultrasonic signals at each time point is obtained i , G represents the ultrasonic signal with a fixed intensity, q represents the number of roof detection points, respectively represent the preset duration from the emission to the reception of the ultrasonic signal and the weight factor of the ultrasonic signal intensity.

[0016] In the second step, the abnormality degree of the roof ultrasonic signals at each time point is compared with the preset abnormality degree threshold of the ultrasonic signal. If the abnormality degree of the roof ultrasonic signal at a certain time point is greater than or equal to the preset abnormality degree threshold of the ultrasonic signal, it is determined that the roof has a cracking situation at that time point. The time point when the abnormality degree of the first roof ultrasonic signal is greater than or equal to the preset abnormality degree threshold of the ultrasonic signal is denoted as the crack start time point.

[0017] Preferably, the specific analysis method for the severity of the cracks in the top plate is as follows: Read the top plate images at each time point, extract the image of the top plate at the crack start time point and each subsequent time point, and re-number them in chronological order as the images at each crack time point, numbered 1, 2,..., z,..., s. Extract the edge contours of each crack from the images at each crack time point through edge detection technology, and obtain the length of each crack in the images at each crack time point, denoted as L zj , where j represents the number of the j-th crack, j = 1, 2,..., g. Through the formula the crack propagation speed ξ is obtained, and L (z-1)j represents the length of the j-th crack in the (z - 1)-th crack time point image, Δt represents the set time interval, s represents the number of crack time points, and g represents the number of cracks. Substitute them into the formula to obtain the severity θ of the cracks in the top plate, where ξ' represents the preset reference value of the crack propagation speed.

[0018] Preferably, the specific detection method for the pressure parameters of the coal seam working face after the roof cutting operation is as follows: First step, select a number of equally spaced monitoring points on the coal seam working face, and detect the pressure at each monitoring point on the coal seam working face before the roof cutting operation through a pressure sensor, denoted as the pressure P at each monitoring point on the coal seam working face f , where f represents the number of the f-th monitoring point, f = 1, 2,..., k. Calculate the average pressure of the coal seam working face by taking the average value of the pressures at each monitoring point on the coal seam working face, denoted as Substitute it into the formula to obtain the pressure uniformity σ of the coal seam working face, where k represents the number of monitoring points.

[0019] Second step, perform the roof cutting operation according to a predetermined plan. After the roof cutting operation is completed, select a number of time points at fixed time intervals, denoted as each monitoring time point, and detect the pressure at each monitoring point on the coal seam working face at each monitoring time point, denoted as the pressure at each monitoring point at each monitoring time point on the coal seam working face. Analyze and obtain the pressure uniformity σ' at each monitoring time point on the coal seam working face according to the method of analyzing the pressure uniformity of the coal seam working face x , where x represents the number of the x-th monitoring time point, x = 1, 2,..., y. At the same time, take adjacent monitoring time points as a group, and obtain the difference in pressure uniformity between each group of adjacent monitoring time points by taking the difference in pressure uniformity between each group of adjacent monitoring time points, and accumulate to obtain the change value of the pressure uniformity of the coal seam working face, denoted as Δσ'.

[0020] Preferably, the specific analysis method for the effect evaluation coefficient of the coal seam roof cutting operation is as follows: Read the pressure uniformity σ of the coal seam working face and the pressure uniformity σ' at each monitoring time point on the coal seam working face respectively x, the change value Δσ' of the pressure uniformity degree of the coal seam working face is substituted into the formula to obtain the effect evaluation coefficient λ of the roof cutting operation of the coal seam, where η1 and η2 respectively represent the preset pressure uniformity degree and the weight factor of the change value of the pressure uniformity degree, and Δσ'0 represents the reference value of the preset change value of the pressure uniformity degree.

[0021] Preferably, the specific detection methods for the roadway deformation amount and the surrounding rock stability are as follows: First step, when the roof collapses, the roadway roof is video-recorded by a video recording device to obtain a roadway roof video. On the roadway roof, several collection points are randomly selected, and the positions of each collection point on the roadway roof in the first frame and the last frame of the roadway roof video are marked respectively to obtain the position points of the first frame and the last frame of each collection point on the roadway roof.

[0022] Second step, the last frame of the roadway roof video is overlapped with the first frame of the roadway roof video, and the vertical distance between the last frame position point and the first frame position point of each collection point on the roadway roof is measured, which is recorded as the subsidence distance of each collection point on the roadway roof, and the roadway deformation amount is analyzed.

[0023] Third step, at the same time, the change of the surrounding rock is video-recorded to obtain a surrounding rock movement video. On the surrounding rock, several position points are randomly selected, and the displacement distance of each position point on the surrounding rock is obtained according to the method of analyzing the subsidence distance of each collection point on the roadway roof, and the surrounding rock stability is analyzed therefrom.

[0024] Preferably, the specific analysis method for the feasibility of the roof cutting and roadway protecting is as follows: The roadway deformation amount ε and the surrounding rock stability σ are read and substituted into the formula to obtain the feasibility χ of the roof cutting and roadway protecting, where a1 and a2 respectively represent the preset weight factors of the roadway deformation amount and the surrounding rock stability, and e represents the natural constant.

[0025] Preferably, the specific analysis method for the comprehensive analysis of the abutment pressure is as follows: The bearing stability ξ of the coal pillar, the severity θ of the roof cracks, the effect evaluation coefficient λ of the roof cutting operation of the coal seam, and the feasibility χ of the roof cutting and roadway protecting are read respectively and substituted into the formula to obtain the influence evaluation index of the abutment pressure on the fully-mechanized caving face of the coal seam w1, w2, w3, and w4 respectively represent the weight factors of the bearing stability of the coal pillar, the severity of the roof cracks, the effect evaluation coefficient of the roof cutting operation of the coal seam, and the feasibility of the roof cutting and roadway protecting. The influence evaluation index of the abutment pressure on the fully-mechanized caving face of the coal seam is compared with the preset threshold value of the influence evaluation index of the abutment pressure. If the influence evaluation index of the abutment pressure on the fully-mechanized caving face of the coal seam is greater than the preset threshold value of the influence evaluation index of the abutment pressure, it indicates that the abutment pressure may cause greater risks to the working face. On the contrary, it indicates that the influence of the abutment pressure on the working face is relatively small, and the working face is in a relatively stable and safe state, and the influence situation of the abutment pressure is fed back.

[0026] Compared with the prior art, the embodiments of the present invention have at least the following advantages or beneficial effects: 1. The present invention collects stress data of the coal pillar, analyzes and obtains the bearing stability of the coal pillar, accurately grasps the state of the coal pillar, ensures its bearing stability, guarantees operation safety, and reduces the risk of collapse.

[0027] 2. The present invention obtains and analyzes the crack propagation speed through images, thereby obtaining the severity of the cracks in the roof, timely understanding the crack conditions in the roof, and making countermeasures in advance to prevent the cracks from further expanding and causing danger.

[0028] 3. The present invention obtains the effect evaluation coefficient of the coal seam top cutting operation according to the pressure parameter analysis of the coal seam working face after the top cutting operation, accurately evaluates the effect of the top cutting operation, helps to optimize the top cutting process and improve the operation quality.

[0029] 4. The present invention obtains the deformation of the tunnel and the stability of the surrounding rock through video recording analysis, and then analyzes the feasibility of cutting the top to protect the tunnel. The deformation of the tunnel and the stability of the surrounding rock are clear, which facilitates the adoption of targeted support measures to maintain the safety of the tunnel.

[0030] 5. The present invention analyzes and obtains the evaluation index of the support pressure influence of the coal seam fully mechanized caving working face according to the bearing stability of the coal pillar, the severity of the cracks in the roof, the effect evaluation coefficient of the coal seam top cutting operation, and the feasibility of top cutting and tunnel protection, which is beneficial to optimize the mining technology and process and improve the mining efficiency. BRIEF DESCRIPTION OF THE DRAWINGS

[0031] The present invention is further described using the accompanying drawings, but the embodiments in the accompanying drawings do not constitute any limitation to the present invention. A person skilled in the art can obtain other drawings based on the following drawings without creative work.

[0032] Figure 1 It is a schematic diagram of the method flow of the present invention.

[0033] Figure 2 for Figure 1 Schematic diagram of the process of tunnel deformation and surrounding rock stability in step S4.

[0034] Figure 3 for Figure 1 Flow judgment block diagram of step S5 in FIG. DETAILED DESCRIPTION

[0035] Next, in combination with the accompanying drawings in the embodiments of the present invention, the technical solutions in the embodiments of the present invention will be clearly and completely described. Obviously, the described embodiments are only a part of the embodiments of the present invention, rather than all the embodiments. Based on the embodiments of the present invention, all other embodiments obtained by those of ordinary skill in the art without creative efforts shall fall within the protection scope of the present invention.

[0036] Please refer to Figure 1 As shown, the present invention provides a simulation analysis method for the abutment pressure distribution in a fully-mechanized caving face of an extra-thick coal seam. The method includes the following steps: S1. Bearing stability analysis: Collect stress data of the coal pillar, and analyze to obtain the bearing stability of the coal pillar. The stress data of the coal pillar includes the proportion of the high-stress area at each time point and the stress value growth rate.

[0037] The specific detection method for the stress data of the coal pillar is as follows: First step, select a number of detection points at different parts of the coal pillar according to the set spacing, and at the same time select a number of time points according to the set time interval. Collect the stress values of each detection point of the coal pillar at each time point through stress sensors, denoted as the stress values of each detection point of the coal pillar at each time point. Compare the stress values of each detection point of the coal pillar at each time point with the preset high-stress value threshold, and screen out each detection point where the stress value of the coal pillar at each time point is greater than or equal to the preset high-stress value threshold, denoted as each high-stress value detection point of the coal pillar at each time point. Connect each high-stress value detection point of the coal pillar at each time point to outline the boundary of the high-stress area of the coal pillar at each time point, and obtain the area of the high-stress area of the coal pillar at each time point, denoted as M i , i represents the number of the i-th time point, i = 1, 2,..., n. Through the formula Obtain the proportion of the high-stress area of the coal pillar at each time point M 煤柱 represents the cross-sectional area of the coal pillar; it can accurately grasp the stress state of different parts of the coal pillar, timely discover the high-stress area, provide a basis for taking targeted reinforcement or prevention measures, and ensure the bearing stability and safety of the coal pillar.

[0038] Second step, obtain the stress value of the coal pillar at each time point by calculating the average value of the stress values of each detection point of the coal pillar at each time point. Take two adjacent time points as a group, denoted as each time point group. Take the difference of the stress values of each time point group of the coal pillar to obtain the stress value difference of each time point group of the coal pillar. Divide the stress value difference of each time point group of the coal pillar by the set time interval to obtain the stress value growth rate of each time point group of the coal pillar, and then obtain the stress value growth rate of the coal pillar through mean calculation, denoted as V; Analyzing the stress value difference and growth rate of each time point group of the coal pillar helps to understand the trend and dynamics of the coal pillar stress change, and provides important information for further mining planning and risk warning.

[0039] The specific analysis method for the bearing stability analysis is as follows: read the proportion of the high-stress area at each time point of the coal pillar respectively the stress value growth rate V of the coal pillar, and substitute it into the formula to obtain the bearing stability ξ of the coal pillar. V' represents the preset reference value of the stress value growth rate, φ1 and φ2 respectively represent the preset weight factors of the proportion of the high-stress area and the stress value growth rate, and n represents the number of time points; potential problems in the bearing stability of the coal pillar can be discovered in time, and countermeasures can be taken in advance to reduce risks and losses.

[0040] It should be noted that in a specific embodiment, φ1 can be set to 0.6 and φ2 can be set to 0.4. The proportion of the high-stress area intuitively reflects the pressure distribution situation borne by the coal pillar. The larger the proportion, the wider the range of the coal pillar bearing high stress, and the greater the impact on its stability. A faster stress value growth rate may mean that the pressure change faced by the coal pillar is more intense, which will pose a greater challenge to its bearing capacity and is more likely to cause stability problems. However, the simple growth rate also needs to be comprehensively judged in combination with factors such as the actual stress value size. Therefore, the weight corresponding to the proportion of the high-stress area is greater.

[0041] S2. Crack analysis: Determine the starting time point of the crack according to the abnormality degree of the roof ultrasonic signal at each time point, and then obtain the crack propagation speed through image acquisition and analysis, so as to obtain the severity of the roof crack.

[0042] The specific analysis method for determining the starting time point of the crack is as follows: First step, obtain the roof images at each time point through a camera. At the same time, select several detection points on the roof at a set distance, denoted as each roof detection point. Apply a coupling agent to each roof detection point on the roof, and emit ultrasonic signals with a fixed intensity to each roof detection point at each time point through an ultrasonic flaw detector, and use a probe to receive the ultrasonic signals reflected from the roof, and record the duration from the emission to the reception of the ultrasonic signals at each roof detection point at each time point, denoted as T im , m represents the number of the mth roof detection point, m = 1, 2,..., q. At the same time, record the intensity of the ultrasonic signals reflected from each roof detection point at each time point as G im , and through the formula obtain the abnormality degree δ of the roof ultrasonic signal at each time point i , G represents the ultrasonic signal with a fixed intensity, q represents the number of roof detection points, respectively represent the preset duration from the emission to the reception of the ultrasonic signal and the weight factor of the ultrasonic signal intensity; Abnormal situations such as possible cracking of the roof can be discovered in time, providing timely information for taking maintenance, reinforcement and other measures to avoid the further expansion of potential safety hazards.

[0043] It should be noted that in a specific embodiment, it can be set to 0.6, it can be set to 0.4. The duration can more directly reflect key information such as the thickness of the roof and the medium conditions. By analyzing the duration, some changes in the roof structure can be more accurately judged, which is of great significance for evaluating the degree of abnormality. Although the ultrasonic signal intensity can also provide certain information, it may be more easily interfered by other factors, such as the performance of the transmitting and receiving devices, environmental noise, etc. Its stability and specificity may be relatively weak. Therefore, the weight corresponding to the duration of the ultrasonic signal from transmission to reception is higher.

[0044] Second, compare the abnormality degree of the roof ultrasonic signal at each time point with the preset abnormality degree threshold of the ultrasonic signal. If the abnormality degree of the roof ultrasonic signal at a certain time point is greater than or equal to the preset abnormality degree threshold of the ultrasonic signal, it is determined that there is a cracking situation on the roof at this time point, and the time point when the abnormality degree of the first roof ultrasonic signal is greater than or equal to the preset abnormality degree threshold of the ultrasonic signal is recorded as the crack start time point; clarifying the crack start time point helps to analyze the process and trend of the roof state change, and provides a strong basis for subsequent maintenance plans and safety management.

[0045] The specific analysis method for the severity of the cracks in the roof is as follows: Read the roof images at each time point, extract the crack start time point and the roof images at each subsequent time point, and re-number them in chronological order as the crack time point images, numbered 1, 2,..., z,..., s. Extract the edge contours of each crack from each crack time point image through edge detection technology, and obtain the length of each crack in each crack time point image, denoted as L zj , j represents the number of the jth crack, j = 1, 2,..., g. Through the formula the crack propagation speed ξ is obtained, L (z-1)j represents the length of the jth crack in the (z - 1)th crack time point image, Δt represents the set time interval, s represents the number of crack time points, g represents the number of cracks, and substitute them into the formula to obtain the severity of the roof cracks θ, ξ' represents the preset reference value of the crack propagation speed; it helps to timely grasp the development dynamics of the cracks, so as to take more targeted countermeasures to ensure operation safety.

[0046] S3. Analysis of the influence of roof cutting operation: According to the pressure parameters of the coal seam working face after the roof cutting operation, the effect evaluation coefficient of the coal seam roof cutting operation is obtained. The pressure parameters include the pressure uniformity and the change value of the pressure uniformity.

[0047] The specific detection method for the pressure parameters of the coal seam working face after the roof cutting operation is as follows: In the first step, several equally spaced monitoring points are selected on the coal seam working face. The pressure of each monitoring point on the coal seam working face before the roof cutting operation is detected through a pressure sensor and denoted as the pressure P of each monitoring point on the coal seam working face f , where f represents the number of the f-th monitoring point, f = 1, 2,..., k. The average pressure of the coal seam working face is obtained by calculating the average value of the pressures of each monitoring point on the coal seam working face and denoted as Substitute it into the formula to obtain the pressure uniformity σ of the coal seam working face. k represents the number of monitoring points; this helps to scientifically evaluate the influence of the roof cutting operation on the pressure distribution of the coal seam working face and provides data support for subsequent construction and safety guarantee

[0048] In the second step, the roof cutting operation is carried out according to a predetermined plan. After the roof cutting operation is completed, several time points are selected at fixed time intervals and denoted as each monitoring time point. The pressure of each monitoring point on the coal seam working face at each monitoring time point is detected and denoted as the pressure of each monitoring point at each monitoring time point on the coal seam working face. The pressure uniformity of the coal seam working face at each monitoring time point is analyzed according to the method of analyzing the pressure uniformity of the coal seam working face and denoted as σ' x , where x represents the number of the x-th monitoring time point, x = 1, 2,..., y. At the same time, taking adjacent monitoring time points as a group, the difference in pressure uniformity between each group of adjacent monitoring time points is obtained by taking the difference in pressure uniformity between each group of adjacent monitoring time points, and the cumulative value is the change value of the pressure uniformity of the coal seam working face, denoted as Δσ'; clarifying the change value of the pressure uniformity can better grasp the pressure change trend and amplitude caused by roof cutting, provide a basis for optimizing the roof cutting plan and adjusting the construction strategy, and improve the construction efficiency and safety

[0049] The specific analysis method for the effect evaluation coefficient of the coal seam roof cutting operation is as follows: Read the pressure uniformity σ of the coal seam working face, the pressure uniformity σ' of each monitoring time point on the coal seam working face x , and the change value Δσ' of the pressure uniformity of the coal seam working face respectively. Substitute them into the formula to obtain the effect evaluation coefficient λ of the coal seam roof cutting operation, where η1 and η2 respectively represent the preset weight factors of the pressure uniformity and the change value of the pressure uniformity, and Δσ'0 represents the preset reference value of the change value of the pressure uniformity; this helps to optimize and improve the roof cutting operation to enhance the safety and efficiency of coal seam mining

[0050] It should be noted that in a specific embodiment, η1 can be set to 0.7 and η2 can be set to 0.3. The pressure uniformity directly reflects the rationality of the pressure distribution on the entire coal seam. If the pressure distribution is uneven, it may lead to local stress concentration, increasing the risk of coal seam instability and having an important impact on the roof cutting effect. The change value of the pressure uniformity mainly reflects the dynamic change of the pressure uniformity over time or the operation process. Although it can also reflect some trends and potential problems, relatively speaking, the initial pressure uniformity state is more critical. Therefore, the weight corresponding to the pressure uniformity is higher.

[0051] S4. Feasibility assessment of roof cutting for roadway protection: When the roof collapses, the roadway deformation and surrounding rock stability are obtained through video recording and analysis, and then the feasibility of roof cutting for roadway protection is analyzed.

[0052] Please refer to Figure 2 As shown, the specific detection methods for the roadway deformation and surrounding rock stability are as follows: First step, when the roof collapses, the roadway roof is video-recorded by a video recording device to obtain a roadway roof video. On the roadway roof, several collection points are randomly selected, and the positions of each collection point on the roadway roof in the first frame and the last frame of the roadway roof video are marked respectively to obtain the position points of the first frame and the last frame of each collection point on the roadway roof; this helps to deeply understand the specific details and dynamic changes of roof collapse and better master the behavior characteristics of the roof.

[0053] Second step, the last frame of the roadway roof video is overlapped with the first frame of the roadway roof video, and the vertical distance between the last frame position point and the first frame position point of each collection point on the roadway roof is measured, denoted as the subsidence distance of each collection point on the roadway roof, and the roadway deformation is analyzed; this provides a quantitative index for evaluating the stability and safety of the roadway, timely discovers potential risks, provides a scientific basis for roadway support design and maintenance, and improves the pertinence and effectiveness of support.

[0054] It should be noted that the specific analysis method for the roadway deformation is: read the subsidence distance of each collection point on the roadway roof, denoted as d c , c represents the number of the c-th collection point, c = 1, 2,..., o, and substitute it into the formula to obtain the roadway deformation ε, where d0 represents the reference subsidence distance of the preset collection point, and o represents the number of position points.

[0055] Third step, at the same time, the change of the surrounding rock is video-recorded to obtain a surrounding rock movement video. On the surrounding rock, several position points are randomly selected, and the displacement distance of each position point on the surrounding rock is obtained according to the method of analyzing the subsidence distance of each collection point on the roadway roof, and then the surrounding rock stability is analyzed; this helps to timely discover the unstable areas and potential dangerous points of the surrounding rock, so as to take targeted reinforcement or prevention measures.

[0056] It should be noted that the specific analysis method for the surrounding rock stability is as follows: read the displacement distances of each position point of the surrounding rock, denoted as d' p , where p represents the serial number of the p-th position point, p = 1, 2,..., l. The average displacement distance of the surrounding rock is obtained by calculating the average value of the displacement distances of each position point of the surrounding rock, denoted as Substitute it into the formula to obtain the surrounding rock stability σ, where l represents the number of position points.

[0057] The specific analysis method for the feasibility of roof cutting and roadway protection is as follows: read the roadway deformation ε and the surrounding rock stability σ, and substitute them into the formula to obtain the feasibility χ of roof cutting and roadway protection, where a1 and a2 respectively represent the weight factors of the preset roadway deformation and the surrounding rock stability, and e represents the natural constant; it can more accurately measure the feasibility of roof cutting and roadway protection, reduce errors and uncertainties, make the engineering plan more reasonable and reliable, and at the same time can predict potential risks and problems in advance, adjust strategies or take countermeasures in a timely manner, and reduce engineering risks and costs.

[0058] It should be noted that in a specific embodiment, a1 can be set to 0.5 and a2 can be set to 0.5. The roadway deformation intuitively reflects the actual changes of the roadway during and after roof cutting. If the deformation is too large, it will seriously affect the normal use and safety of the roadway, which is directly related to the feasibility of the roof cutting and roadway protection plan. The surrounding rock stability is an important basis for the long-term stability of the roadway. If the surrounding rock is unstable, even if the current roadway deformation is small, larger problems may occur in the future, which has a profound impact on the long-term effect and feasibility of roof cutting and roadway protection. Therefore, the weights corresponding to the roadway deformation and the surrounding rock stability are equal.

[0059] S5. Comprehensive analysis of abutment pressure: According to the bearing stability ξ of the coal pillar, the severity θ of the roof cracks, the effect evaluation coefficient λ of the coal seam roof cutting operation, and the feasibility χ of the roof cutting and roadway protection, analyze and obtain the influence evaluation index of the abutment pressure in the fully-mechanized caving face of the coal seam and give feedback.

[0060] Please refer to Figure 3 As shown, the specific analysis method for the comprehensive analysis of the abutment pressure is as follows: respectively read the bearing stability ξ of the coal pillar, the severity θ of the roof cracks, the effect evaluation coefficient λ of the coal seam roof cutting operation, and the feasibility χ of the roof cutting and roadway protection, and substitute them into the formula to obtain the influence evaluation index of the abutment pressure in the fully-mechanized caving face of the coal seam w1, w2, w3, and w4 respectively represent the weight factors of the bearing stability of the coal pillar, the severity of roof cracks, the evaluation coefficient of the effect of roof cutting operation in the coal seam, and the feasibility of roadway protection by roof cutting. The evaluation index of the influence of abutment pressure in the fully-mechanized caving face of the coal seam is compared with the preset threshold value of the evaluation index of abutment pressure influence. If the evaluation index of the influence of abutment pressure in the fully-mechanized caving face of the coal seam is greater than the preset threshold value of the evaluation index of abutment pressure influence, it indicates that the abutment pressure may pose a greater risk to the working face. On the contrary, it indicates that the influence of abutment pressure on the working face is relatively small, and the working face is in a relatively stable and safe state, and the influence of abutment pressure is fed back. It can clearly judge the degree of influence of abutment pressure on the working face and give timely feedback, which helps to take targeted measures in advance to cope with possible risks, such as strengthening support, adjusting mining technology, etc., so as to ensure the safe and stable production of the working face.

[0061] It should be noted that in a specific embodiment, w1 can be set to 0.3, w2 can be set to 0.3, w3 can be set to 0.3, and w4 can be set to 0.1. The bearing stability of the coal pillar is directly related to whether the support structure of the entire working face is stable, and has a key influence on the distribution and magnitude of abutment pressure. If the coal pillar is unstable, it may lead to abnormal changes in abutment pressure and affect the safety of the entire working face. The severity of roof cracks will affect the integrity and bearing capacity of the roof, and further affect the dispersion and transmission of abutment pressure. Severe cracks may lead to unreasonable distribution of abutment pressure and increase potential risks. The effect of roof cutting operation will directly change the distribution state of abutment pressure. If the roof cutting effect is good, it can effectively reduce problems such as concentration of abutment pressure, which is very important for the evaluation of abutment pressure influence. Roadway protection by roof cutting mainly considers the influence of abutment pressure from a long-term and overall perspective, but relatively speaking, its direct influence may not be as urgent and critical as the previous ones. Therefore, the bearing stability of the coal pillar, the severity of roof cracks, and the evaluation coefficient of the effect of roof cutting operation in the coal seam have higher corresponding weights.

[0062] The present invention collects stress data of the coal pillar, analyzes to obtain the bearing stability of the coal pillar, obtains the severity of roof cracks by analyzing the crack propagation speed through images, analyzes to obtain the evaluation coefficient of the effect of roof cutting operation in the coal seam according to the pressure parameters of the coal seam working face after roof cutting operation, analyzes to obtain the deformation amount of the roadway and the stability of surrounding rocks by video recording analysis, and then analyzes to obtain the feasibility of roadway protection by roof cutting, so as to comprehensively analyze the evaluation index of the influence of abutment pressure in the fully-mechanized caving face of the coal seam, can better take targeted measures to ensure operation safety, is beneficial to optimizing mining technology and process, and improving mining efficiency.

[0063] Although the embodiments of the present invention have been shown and described above, it can be understood that the above embodiments are exemplary and should not be construed as limiting the present invention. Those of ordinary skill in the art can make changes, modifications, substitutions, and variations to the above embodiments within the scope of the present invention, and still be covered by the protection scope of the present invention.

Claims

1. A simulation analysis method for the abutment pressure distribution in a fully-mechanized caving face of an extra-thick coal seam, characterized in that It includes the following steps: S1. Bearing stability analysis: Collect stress data of the coal pillar, and analyze to obtain the bearing stability of the coal pillar. The stress data of the coal pillar includes the proportion of the high-stress area at each time point and the stress value growth rate; S2. Crack analysis: Determine the crack start time point according to the abnormal degree of the roof ultrasonic signal at each time point, and then obtain the crack propagation speed through image acquisition and analysis, so as to obtain the severity of the roof cracks; S3. Analysis of the influence of roof cutting operation: Analyze to obtain the effect evaluation coefficient of the coal seam roof cutting operation according to the pressure parameters of the coal seam working face after the roof cutting operation. The pressure parameters include the pressure uniformity and the change value of the pressure uniformity; S4. Feasibility evaluation of roof cutting for roadway protection: When the roof collapses, analyze the roadway deformation and surrounding rock stability through video recording, and then analyze the feasibility of roof cutting for roadway protection; S5. Comprehensive analysis of abutment pressure: Based on the bearing stability ξ of the coal pillar, the severity θ of roof cracks, the effect evaluation coefficient λ of roof cutting operation in the coal seam, and the feasibility χ of roadway protection by roof cutting, the influence evaluation index of abutment pressure in the fully-mechanized caving face of the coal seam is analyzed And feedback is carried out.

2. A simulation analysis method for the abutment pressure distribution in a fully-mechanized caving face of an extra-thick coal seam according to claim 1, characterized in that: The specific detection method of the stress data of the coal pillar is as follows: First step: Select a number of detection points at different parts of the coal pillar according to the set spacing, and at the same time select a number of time points according to the set time interval. Collect the stress values of each detection point of the coal pillar at each time point through stress sensors, denoted as the stress values of each detection point of the coal pillar at each time point. Compare the stress values of each detection point of the coal pillar at each time point with the preset high stress value threshold, and screen out each detection point where the stress value of the coal pillar at each time point is greater than or equal to the preset high stress value threshold, denoted as each high stress value detection point of the coal pillar at each time point. Connect each high stress value detection point of the coal pillar at each time point to outline the boundary of the high stress area of the coal pillar at each time point, and obtain the area of the high stress area of the coal pillar at each time point, denoted as M i , where i represents the number of the i-th time point, i = 1, 2,..., n, and through the formula obtain the proportion of the high stress area of the coal pillar at each time point M 煤柱 represents the cross-sectional area of the coal pillar; In the second step, calculate the average value of the stress values of each detection point of the coal pillar at each time point to obtain the stress value of the coal pillar at each time point. Take two adjacent time points as a group, denoted as each time point group. Take the difference of the stress values of each time point group of the coal pillar to obtain the stress value difference of each time point group of the coal pillar. Divide the stress value difference of each time point group of the coal pillar by the set time interval to obtain the stress value growth rate of each time point group of the coal pillar, and then calculate the stress value growth rate of the coal pillar through average value calculation, denoted as V.

3. A simulation analysis method for the abutment pressure distribution in a fully-mechanized caving face of an extra-thick coal seam according to claim 2, characterized in that: The specific analysis method of the bearing stability analysis is as follows: Read the proportion of the high-stress area at each time point of the coal pillar respectively The stress value growth rate V of the coal pillar is substituted into the formula The bearing stability ξ of the coal pillar is obtained, where V' represents the reference value of the preset stress value growth rate, φ1 and φ2 respectively represent the weight factors of the proportion of the preset high-stress area and the stress value growth rate, and n represents the number of time points.

4. A simulation analysis method for the abutment pressure distribution in a fully-mechanized caving face of an extra-thick coal seam according to claim 3, characterized in that: The specific analysis method of determining the crack start time point is as follows: First step: Obtain the roof images at each time point through a camera. Meanwhile, select a number of detection points on the roof at a set distance, denoted as each roof detection point. Apply a coupling agent to each roof detection point on the roof, and transmit ultrasonic signals with a fixed intensity to each roof detection point at each time point through an ultrasonic flaw detector. Use a probe to receive the ultrasonic signals reflected from the roof, and record the duration from the emission to the reception of the ultrasonic signals at each roof detection point at each time point, denoted as T im , where m represents the number of the m-th roof detection point, m = 1, 2,..., q. Meanwhile, denote the intensity of the ultrasonic signals reflected from each roof detection point at each time point as G im , and obtain the abnormality degree δ of the roof ultrasonic signals at each time point through the formula . Here, G represents the ultrasonic signal with a fixed intensity, q represents the number of roof detection points, i , respectively represent the preset duration from the emission to the reception of the ultrasonic signal and the weight factor of the ultrasonic signal intensity; ​ In the second step, compare the abnormal degree of the roof ultrasonic signal at each time point with the preset abnormal degree threshold of the ultrasonic signal. If the abnormal degree of the roof ultrasonic signal at a certain time point is greater than or equal to the preset abnormal degree threshold of the ultrasonic signal, it is judged that there is a cracking situation on the roof at this time point. Record the time point when the abnormal degree of the first roof ultrasonic signal is greater than or equal to the preset abnormal degree threshold of the ultrasonic signal as the crack start time point.

5. A simulation analysis method for the abutment pressure distribution in a fully-mechanized caving face of an extra-thick coal seam according to claim 4, characterized in that: The specific analysis method of the severity of the roof cracks is as follows: Read the roof images at each time point, extract the roof images at the crack start time point and subsequent time points, and renumber them in chronological order as crack time point images, numbered 1, 2,..., z,..., s. Extract the edge contours of each crack from the crack time point images through edge detection technology, and obtain the lengths of each crack in the crack time point images, denoted as L zj , where j represents the number of the j-th crack, j = 1, 2,..g., through the formula obtain the crack propagation speed ξ, L (z-1)j represents the length of the j-th crack in the (z - 1)-th crack time point image, Δt represents the set time interval, s represents the number of crack time points, g represents the number of cracks, and substitute them into the formula obtain the crack severity θ of the roof, where ξ' represents the preset reference value of the crack propagation speed.

6. A simulation analysis method for the abutment pressure distribution in a fully-mechanized caving face of an extra-thick coal seam according to claim 1, characterized in that: The specific detection method of the pressure parameters of the coal seam working face after the roof cutting operation is as follows: First step, select a number of equally spaced monitoring points on the coal seam working face, and detect the pressure of each monitoring point on the coal seam working face before the roof cutting operation through a pressure sensor, denoted as the pressure P of each monitoring point on the coal seam working face f , f represents the number of the f-th monitoring point, f = 1, 2,..., k. The average pressure of the coal seam working face is obtained by taking the average value of the pressures of each monitoring point on the coal seam working face, denoted as Substitute it into the formula to obtain the pressure uniformity σ of the coal seam working face, where k represents the number of monitoring points; In the second step, roof cutting operation is carried out according to the predetermined plan. After the roof cutting operation is completed, several time points are selected at fixed time intervals and recorded as each monitoring time point. The pressure of each monitoring point on the coal seam working face at each monitoring time point is detected and recorded as the pressure of each monitoring point at each monitoring time point on the coal seam working face. The pressure uniformity degree of the coal seam working face at each monitoring time point is analyzed by the method of analyzing the pressure uniformity degree of the coal seam working face and is recorded as σ'. x , where x represents the number of the x-th monitoring time point, x = 1, 2,..., y. At the same time, taking adjacent monitoring time points as a group, the difference in the pressure uniformity degree of each group of adjacent monitoring time points is obtained by taking the difference in the pressure uniformity degree of each group of adjacent monitoring time points, and the change value of the pressure uniformity degree of the coal seam working face is accumulated and recorded as Δσ'.

7. A method for simulating and analyzing the abutment pressure distribution in a fully-mechanized caving face of an extra-thick coal seam according to claim 6, characterized in that: The specific analysis method of the effect evaluation coefficient of the coal seam roof cutting operation is as follows: Read the pressure uniformity degree σ of the coal seam working face and the pressure uniformity degree σ' at each monitoring time point of the coal seam working face respectively x , the pressure uniformity degree change value Δσ' of the coal seam working face, and substitute them into the formula to obtain the effect evaluation coefficient λ of the roof cutting operation of the coal seam, where η1 and η2 respectively represent the preset weight factors of the pressure uniformity degree and the pressure uniformity degree change value, and Δσ'0 represents the reference value of the preset pressure uniformity degree change value.

8. A method for simulating and analyzing the abutment pressure distribution in a fully-mechanized caving face of an extra-thick coal seam according to claim 1, characterized in that: The specific detection method of the roadway deformation and surrounding rock stability is as follows: In the first step, when the roof collapses, use a video recording device to record the video of the roadway roof to obtain the roadway roof video. Select several acquisition points on the roadway roof, and mark the positions of each acquisition point on the roadway roof in the first frame and the last frame of the roadway roof video respectively to obtain the position points of the first frame and the last frame of each acquisition point on the roadway roof; In the second step, overlap the last frame of the roadway roof video with the first frame of the roadway roof video, measure the vertical distance between the last frame position point and the first frame position point of each acquisition point on the roadway roof, denoted as the subsidence distance of each acquisition point on the roadway roof, and analyze to obtain the roadway deformation. In the third step, simultaneously record the changes in the surrounding rock by video to obtain a video of the surrounding rock movement. Select several position points on the surrounding rock, and obtain the displacement distances of each position point of the surrounding rock according to the method of analyzing the subsidence distance of each collection point on the roadway roof, and thus analyze the stability of the surrounding rock.

9. A simulation analysis method for the abutment pressure distribution in a fully-mechanized caving face of an extra-thick coal seam according to claim 8, characterized in that: The specific analysis method for the feasibility of the roof cutting and roadway protection is as follows: Read the roadway deformation ε and the surrounding rock stability σ, and substitute them into the formula respectively to obtain the feasibility χ of roof cutting for roadway protection, where a1 and a2 respectively represent the preset weight factors of roadway deformation and surrounding rock stability, and e represents the natural constant.

10. A method for simulating and analyzing the abutment pressure distribution in a fully-mechanized caving face of an extra-thick coal seam according to claim 1, characterized in that: The specific analysis method for the comprehensive analysis of abutment pressure is as follows: Read the bearing stability ξ of the coal pillar, the severity θ of the roof cracks, the effect evaluation coefficient λ of the coal seam roof cutting operation, and the feasibility χ of the roof cutting roadway protection respectively, and substitute them into the formula Obtain the evaluation index of the influence of the abutment pressure on the fully-mechanized caving face of the coal seam w1, w2, w3, and w4 respectively represent the weight factors of the bearing stability of the coal pillar, the severity of the roof cracks, the effect evaluation coefficient of the coal seam roof cutting operation, and the feasibility of the roof cutting roadway protection. Compare the evaluation index of the influence of the abutment pressure on the fully-mechanized caving face of the coal seam with the preset threshold of the evaluation index of the influence of the abutment pressure. If the evaluation index of the influence of the abutment pressure on the fully-mechanized caving face of the coal seam is greater than the preset threshold of the evaluation index of the influence of the abutment pressure, it indicates that the abutment pressure may pose a greater risk to the working face. Otherwise, it indicates that the influence of the abutment pressure on the working face is relatively small, and the working face is in a relatively stable and safe state, and feedback on the influence of the abutment pressure is carried out.

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