A Simulation and Analysis Method for Support Pressure Distribution in Fully Mechanized Longwall Mining Faces of Extra-Thick Coal Seams

Abstract

This invention relates to the field of mining engineering technology, specifically to a method for simulating and analyzing the support pressure distribution in a fully mechanized longwall mining face of an extra-thick coal seam. The method includes the following steps: S1. Bearing stability analysis; S2. Crack analysis; S3. Analysis of the impact of roof cutting operations; S4. Feasibility assessment of roof cutting and roadway protection; S5. Comprehensive analysis of support pressure. This invention collects stress data from coal pillars to analyze their bearing stability. It acquires and analyzes crack propagation rates through image processing to determine the severity of roof cracks. Based on the pressure parameters of the coal seam working face after roof cutting, it analyzes the effectiveness evaluation coefficient of the roof cutting operation. Through video recording and analysis, it obtains roadway deformation and surrounding rock stability, thereby analyzing the feasibility of roof cutting and roadway protection. This comprehensive analysis of the support pressure impact evaluation index of the fully mechanized longwall mining face allows for better implementation of targeted measures to ensure operational safety, optimizes mining processes and procedures, and improves mining efficiency.

Description

Technical Field

[0001] This invention relates to the field of mining engineering technology, and more specifically, to a method for simulating and analyzing the support pressure distribution in a fully mechanized longwall mining face of an extra-thick coal seam. Background Technology

[0002] In the coal mining industry, fully mechanized longwall mining is an important method. With the continuous development of coal resource extraction, higher demands are being placed on the safety and efficiency of the longwall face.

[0003] Understanding the distribution of bearing pressure in fully mechanized longwall mining faces is crucial for ensuring mining safety and optimizing mining processes. The distribution of bearing pressure directly affects key aspects such as the stability of the surrounding rock, the effectiveness of roof support, and the advancement of the working face. Through simulation studies of bearing pressure distribution, we can deeply analyze the pressure variation patterns under different mining conditions, providing a scientific basis for the rational design of roadway layout, selection of support methods, and determination of mining parameters.

[0004] For example, Chinese Patent Application No. 201810751300.1 discloses a method for determining the spacing of pressure relief boreholes based on target support pressure. This method measures the stress distribution around the borehole by measuring the original rock stress and the curve and peak value of the advanced support pressure in the coal seam and calculating the radius of the plastic zone of the borehole. Then, pressure relief boreholes with increasing spacing are constructed along the mining direction of the working face, and borehole stress gauges are installed in the boreholes to monitor the changes in borehole stress in real time as the working face advances. This allows for the direct acquisition of the support pressure at different borehole locations, enabling a relatively accurate understanding of the stress situation in the coal seam and the stress distribution around the borehole.

[0005] However, the above-mentioned patent has the following problems: the solution mainly focuses on the measurement and analysis of stress distribution around the borehole. However, this is not comprehensive enough. The analysis of coal seams requires detailed analysis and evaluation from multiple aspects, such as coal pillar bearing capacity, roof cracks, and roof cutting operations. If a comprehensive evaluation is not performed, it may lead to an inaccurate grasp of the overall condition of the coal seam, which may result in a series of problems such as unreasonable mining planning, increased safety risks, and low mining efficiency, which seriously affect the smooth progress and benefits of coal seam mining. Summary of the Invention

[0006] To overcome the shortcomings in the prior art, this invention provides a method for simulating and analyzing the support pressure distribution in a fully mechanized longwall mining face of an extra-thick coal seam, which can effectively solve the problems mentioned in the prior art.

[0007] The objective of this invention can be achieved through the following technical solution: a method for simulating and analyzing the bearing pressure distribution of an extra-thick coal seam fully mechanized caving face, the method comprising the following steps: S1. Bearing stability analysis: collecting stress data of the coal pillar and analyzing the bearing stability of the coal pillar, the stress data of the coal pillar including the proportion of high stress areas at each time point and the stress value growth rate.

[0008] S2. Crack Analysis: The crack initiation time is determined based on the degree of abnormality of the ultrasonic signal of the roof at each time point. Then, the crack propagation rate is analyzed by acquiring images to obtain the severity of the crack in the roof.

[0009] S3. Impact analysis of roof cutting operation: The effect evaluation coefficient of the roof cutting operation is obtained by analyzing 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.

[0010] S4. Feasibility assessment of roof cutting and tunnel protection: When the roof collapses, the deformation of the tunnel and the stability of the surrounding rock are obtained by video recording and analysis, and then the feasibility of roof cutting and tunnel protection is analyzed.

[0011] S5. Comprehensive Analysis of Support Pressure: Based on the bearing stability ξ of the coal pillar, the severity of roof cracks θ, the evaluation coefficient λ of the effect of coal seam roof cutting operation, and the feasibility χ of roof cutting and roadway protection, the influence evaluation index of support pressure on the coal seam fully mechanized longwall face is obtained. And provide feedback.

[0012] Preferably, the specific method for detecting the stress data of the coal pillar is as follows: First, select several detection points at different locations of the coal pillar according to a set interval, and simultaneously select several time points according to a set time interval. Collect the stress values ​​of each detection point of the coal pillar at each time point using a stress sensor, and record them as the stress values ​​of each detection point at each time point of the coal pillar. Compare the stress values ​​of each detection point at each time point of the coal pillar with a preset high stress value threshold, and select the detection points whose stress values ​​at each time point of the coal pillar are greater than or equal to the preset high stress value threshold, and record them as the high stress value detection points of each time point of the coal pillar. By connecting the high stress value detection points of each time point of the coal pillar, the boundary of the high stress region of each time point of the coal pillar is delineated, and the area of ​​the high stress region of each time point of the coal pillar is obtained, denoted as M. i Let i represent the number of the i-th time point, i = 1, 2, ..., n, obtained through the formula... The proportion of high-stress areas at each time point of the coal pillar was obtained. M 煤柱 This represents the cross-sectional area of ​​the coal column.

[0013] The second step involves averaging the stress values ​​at each detection point of the coal pillar at each time point to obtain the stress values ​​at each time point. Two adjacent time points are grouped together and denoted as each time point group. The stress values ​​of each time point group are subtracted to obtain the stress value difference of each time point group. The stress value growth rate of each time point group is obtained by dividing the stress value difference of each time point group by a set time interval. The average value is then used to calculate the stress value growth rate of the coal pillar, denoted as V.

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

[0015] Preferably, the specific analysis method for determining the crack initiation time point is as follows: First, acquire images of the top plate at various time points using a camera, and simultaneously select several detection points on the top plate at a set distance, denoted as each top plate detection point. Apply coupling agent to each top plate detection point, and use an ultrasonic flaw detector to emit ultrasonic signals of fixed intensity to each top plate detection point at each time point. Receive the ultrasonic signals reflected back from the top plate using a probe, and record the time from emission to reception of the ultrasonic signals at each top plate detection point at each time point, denoted as T. im Let m represent the number of the m-th roof detection point, m = 1, 2, ..., q. The intensity of the ultrasonic signal reflected from each roof detection point at each time point is denoted as G. im Through formula The degree of anomaly δ of the ultrasonic signal of the roof at each time point was obtained. i G represents a fixed-intensity ultrasonic signal, and q represents the number of detection points on the top plate. These represent the preset duration of the ultrasonic signal from transmission to reception and the weighting factor of the ultrasonic signal intensity, respectively.

[0016] The second step is to compare the degree of abnormality of the ultrasonic signal of the roof at each time point with the preset threshold for the degree of abnormality of the ultrasonic signal. If the degree of abnormality of the ultrasonic signal of the roof at a certain time point is greater than or equal to the preset threshold for the degree of abnormality of the ultrasonic signal, it is determined that there is a crack in the roof at that time point. The time point when the degree of abnormality of the ultrasonic signal of the roof is greater than or equal to the preset threshold for the degree of abnormality of the ultrasonic signal is recorded as the crack start time point.

[0017] Preferably, the specific method for analyzing the severity of the cracks in the roof slab is as follows: Read roof slab images at various time points, extract the crack initiation time point and subsequent roof slab images from them, and renumber them chronologically as crack time point images, numbered 1, 2, ..., z, ..., s. Extract the edge contours of each crack from each crack time point image using edge detection technology, and obtain the length of each crack in each crack time point image, denoted as L. zj Let j represent the number of the j-th crack, where j = 1, 2, ..., g. This is determined using the formula... The crack propagation rate ξ, L is obtained. (z-1)j Let Δt represent the length of the j-th crack in the image at the (z-1)-th crack time point, Δt represent the set time interval, s represent the number of crack time points, and g represent the number of cracks. Substitute these values ​​into the formula. The severity of the cracks in the top plate is obtained as θ, where ξ' represents the preset reference value for crack propagation rate.

[0018] Preferably, the specific method for detecting the pressure parameters of the coal seam working face after the roof cutting operation is as follows: First, select several equally spaced monitoring points on the coal seam working face, and use pressure sensors to detect the pressure at each monitoring point on the coal seam working face before the roof cutting operation, and record it as the pressure P at each monitoring point on the coal seam working face. f Let f represent the number of the f-th monitoring point, f = 1, 2, ..., k. The average pressure of the coal seam working face is obtained by averaging the pressure at each monitoring point, denoted as . Substitute it into the formula The pressure uniformity σ at the coal seam working face is obtained, where k represents the number of monitoring points.

[0019] The second step involves performing roof cutting according to the predetermined plan. After the roof cutting is completed, several time points are selected at fixed intervals and recorded as monitoring time points. The pressure at each monitoring point on the coal seam face at each monitoring time point is measured and recorded as the pressure at each monitoring point on the coal seam face at each monitoring time point. The pressure uniformity at each monitoring time point on the coal seam face is analyzed using the method for analyzing the pressure uniformity of the coal seam face and recorded as σ'. x x represents the number of the xth monitoring time point, x = 1, 2, ..., y. At the same time, adjacent monitoring time points are grouped together. The pressure uniformity difference between adjacent monitoring time points in each group is obtained by subtracting the pressure uniformity of adjacent monitoring time points in each group. The sum of these values ​​is used to obtain the pressure uniformity change value of the coal seam working face, which is 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' of each monitoring time point of the coal seam working face. xThe variation value of pressure uniformity at the coal seam working face, Δσ', is substituted into the formula. The effect evaluation coefficient λ of the coal seam roof cutting operation is obtained, where η1 and η2 represent the preset pressure uniformity and pressure uniformity variation value weight factors, respectively, and Δσ'0 represents the preset pressure uniformity variation value reference value.

[0021] Preferably, the specific detection method for the roadway deformation and surrounding rock stability is as follows: First, when the roof collapses, the roadway roof is recorded by video recording equipment to obtain a video of the roadway roof. Several collection points are randomly selected on the roadway roof, and the positions of each collection point on the roadway roof in the first and last frames of the roadway roof video are marked to obtain the position points of each collection point on the roadway roof in the first and last frames.

[0022] The second step is to overlay the last frame of the tunnel roof video with the first frame of the tunnel roof video, measure the vertical distance between the last frame position point and the first frame position point of each acquisition point on the tunnel roof, record it as the subsidence distance of each acquisition point on the tunnel roof, and analyze the tunnel deformation.

[0023] The third step involves simultaneously recording video of the changes in the surrounding rock to obtain a video of the movement of the surrounding rock. Several locations on the surrounding rock are randomly selected, and the displacement distance of each location is obtained by analyzing the subsidence distance of each collection point on the roof of the tunnel. The stability of the surrounding rock is then analyzed.

[0024] Preferably, the specific method for analyzing the feasibility of the roof-cutting and tunneling is as follows: read the tunnel deformation ε and the surrounding rock stability σ, and substitute them into the formula respectively. The feasibility χ of roof cutting and tunnel protection is obtained, where a1 and a2 represent the preset weighting factors of tunnel deformation and surrounding rock stability, respectively, and e represents the natural constant.

[0025] Preferably, the specific analysis method for the comprehensive analysis of the support pressure is as follows: The bearing stability ξ of the coal pillar, the severity of cracks in the roof θ, the evaluation coefficient λ of the effect of roof cutting operation, and the feasibility χ of roof cutting and roadway protection are respectively read and substituted into the formula. The evaluation index of the influence of support pressure on the fully mechanized coal seam mining face was obtained. w1, w2, w3, and w4 represent the bearing stability of the coal pillar, the severity of cracks in the roof, the evaluation coefficient of the effect of coal seam roof cutting operation, and the weighting factor of the feasibility of roof cutting and roadway protection, respectively. The evaluation index of the impact of the support pressure on the coal seam fully mechanized caving face is compared with the preset threshold of the support pressure impact evaluation index. If the evaluation index of the impact of the support pressure on the coal seam fully mechanized caving face is greater than the preset threshold of the support pressure impact evaluation index, it indicates that the support pressure may pose a greater risk to the working face. Conversely, it indicates that the impact of the support pressure on the working face is relatively small, and the working face is in a relatively stable and safe state. Feedback is provided on the impact of the support pressure.

[0026] Compared with the prior art, the embodiments of the present invention have at least the following advantages or beneficial effects: First, the present invention collects and analyzes the stress data of the coal pillar to obtain the bearing stability of the coal pillar, accurately grasps the state of the coal pillar, ensures its bearing stability, guarantees operational safety, and reduces risks such as collapse.

[0027] Second, this invention uses image acquisition and analysis to determine the crack propagation rate, thereby obtaining the severity of the cracks in the roof slab, enabling timely understanding of the crack situation, and allowing for proactive countermeasures to prevent further crack expansion and potential dangers.

[0028] Third, this invention obtains the effect evaluation coefficient of the coal seam roof cutting operation based on the pressure parameters of the coal seam working face after the roof cutting operation, accurately evaluates the effect of the roof cutting operation, and helps to optimize the roof cutting process and improve the quality of operation.

[0029] Fourth, this invention obtains the amount of roadway deformation and the stability of the surrounding rock through video recording and analysis, and then analyzes the feasibility of roof cutting and roadway protection. It clearly understands the roadway deformation and the stability of the surrounding rock, which facilitates the implementation of targeted support measures to maintain roadway safety.

[0030] Fifth, based on the bearing stability of the coal pillar, the severity of cracks in the roof, the evaluation coefficient of the effect of coal seam roof cutting operation, and the feasibility of roof cutting and roadway protection, this invention analyzes and obtains the evaluation index of the influence of the support pressure on the fully mechanized longwall mining face, which is conducive to optimizing the mining technology and process and improving mining efficiency. Attached Figure Description

[0031] The present invention will be further described with reference to the accompanying drawings, but the embodiments in the drawings do not constitute any limitation on the present invention. For those skilled in the art, other drawings can be obtained based on the following drawings without creative effort.

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

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

[0034] Figure 3 for Figure 1 The flowchart for step S5. Detailed Implementation

[0035] The technical solutions of the embodiments of the present invention will be clearly and completely described below with reference to the accompanying drawings. Obviously, the described embodiments are only some embodiments of the present invention, and not all embodiments. Based on the embodiments of the present invention, all other embodiments obtained by those skilled in the art without creative effort are within the scope of protection of the present invention.

[0036] Please see Figure 1 As shown, the present invention provides a method for simulating and analyzing the bearing pressure distribution in a fully mechanized longwall mining face of an extra-thick coal seam. The method includes the following steps: S1. Bearing stability analysis: The stress data of the coal pillar is collected and analyzed to obtain the bearing stability of the coal pillar. The stress data of the coal pillar includes the proportion of high stress areas at each time point and the stress value growth rate.

[0037] The specific method for detecting the stress data of the coal pillar is as follows: First, select several detection points at different locations of the coal pillar according to a set interval, and simultaneously select several time points according to a set time interval. Collect the stress values ​​of each detection point at each time point using a stress sensor, and record them as the stress values ​​of each detection point at each time point of the coal pillar. Compare the stress values ​​of each detection point at each time point of the coal pillar with a preset high stress threshold, and select the detection points whose stress values ​​at each time point of the coal pillar are greater than or equal to the preset high stress threshold, and record them as the high stress value detection points of each time point of the coal pillar. Connect the high stress value detection points of each time point of the coal pillar to delineate the boundaries of the high stress regions of each time point of the coal pillar, and obtain the area of ​​the high stress regions of each time point of the coal pillar, denoted as M. i Let i represent the number of the i-th time point, i = 1, 2, ..., n, obtained through the formula... The proportion of high-stress areas at each time point of the coal pillar was obtained. M 煤柱 It represents the cross-sectional area of ​​the coal pillar; it can accurately grasp the stress state of different parts of the coal pillar, promptly identify high-stress areas, provide a basis for taking targeted reinforcement or prevention measures, and ensure the load-bearing stability and safety of the coal pillar.

[0038] The second step involves averaging the stress values ​​at each monitoring point within the coal pillar at various time points to obtain the stress values ​​at each time point. Two adjacent time points are grouped together, and the stress values ​​of each time point group are subtracted to obtain the stress value difference. The stress value growth rate of each time point group is obtained by dividing the stress value difference by a set time interval. The average value is then used to calculate the stress value growth rate of the coal pillar, denoted as V. Analyzing the stress value difference and growth rate of each time point group helps to understand the trend and dynamics of stress changes in the coal pillar, providing important information for further mining planning and risk warning.

[0039] The specific analysis method for the bearing stability analysis is as follows: The proportion of high-stress areas in the coal pillar is read at each time point. The rate of increase V of stress in the coal pillar is substituted into the formula. The bearing stability ξ of the coal pillar is obtained, where V' represents the preset reference value of stress value growth rate, φ1 and φ2 represent the preset proportion of high stress area and the weighting factor of stress value growth rate, respectively, and n represents the number of time points; this allows for the timely detection of potential problems in the bearing stability of the coal pillar, enabling proactive countermeasures to reduce risks and losses.

[0040] It should be noted that, in one specific embodiment, φ1 can be set to 0.6 and φ2 can be set to 0.4. The proportion of high-stress areas directly reflects the pressure distribution of the coal pillar. The larger the proportion, the wider the range of high stress the coal pillar bears, 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 drastic, which will pose a greater challenge to its bearing capacity and is more likely to cause stability problems. However, the growth rate alone needs to be judged in combination with factors such as the actual stress value. Therefore, the proportion of high-stress areas has a greater weight.

[0041] S2. Crack Analysis: The crack initiation time is determined based on the degree of abnormality of the ultrasonic signal of the roof at each time point. Then, the crack propagation rate is analyzed by acquiring images to obtain the severity of the crack in the roof.

[0042] The specific analytical method for determining the crack initiation time is as follows: First, acquire images of the top plate at various time points using a camera. Simultaneously, select several detection points on the top plate at a set distance, denoted as each top plate detection point. Apply coupling agent to each top plate detection point. Then, use an ultrasonic flaw detector to emit ultrasonic signals of fixed intensity to each top plate detection point at each time point. Receive the ultrasonic signals reflected back from the top plate using a probe. Record the time from emission to reception of the ultrasonic signals at each top plate detection point at each time point, denoted as T. im Let m represent the number of the m-th roof detection point, m = 1, 2, ..., q. The intensity of the ultrasonic signal reflected from each roof detection point at each time point is denoted as G. im Through formula The degree of anomaly δ of the ultrasonic signal of the roof at each time point was obtained. i G represents a fixed-intensity ultrasonic signal, and q represents the number of detection points on the top plate. These represent the preset duration of the ultrasonic signal from transmission to reception and the weighting factor of the ultrasonic signal intensity, respectively; they can promptly detect abnormalities such as cracks in the roof, providing timely information for maintenance and reinforcement measures, and preventing further expansion of safety hazards.

[0043] It should be noted that, in one specific embodiment, It can be set to 0.6. The duration can be set to 0.4. The duration can directly reflect key information such as the thickness of the top plate and the condition of the medium. By analyzing the duration, some changes in the top plate structure can be judged more accurately, which is of great significance for assessing the degree of anomaly. Although the ultrasonic signal intensity can also provide some information, it may be more susceptible to interference from other factors, such as the performance of the transmitting and receiving devices and environmental noise. Its stability and specificity may be relatively weak. Therefore, the duration of the ultrasonic signal from transmission to reception has a higher weight.

[0044] The second step involves comparing the degree of abnormality of the ultrasonic signals from the roof at each time point with a preset threshold for the degree of abnormality of the ultrasonic signals. If the degree of abnormality of the ultrasonic signals from the roof at a certain time point is greater than or equal to the preset threshold, it is determined that there is a crack in the roof at that time point. The time point when the degree of abnormality of the first ultrasonic signal from the roof is greater than or equal to the preset threshold is recorded as the crack initiation time point. Determining the crack initiation time point helps to analyze the process and trend of roof condition changes, providing a strong basis for subsequent maintenance plans and safety management.

[0045] The specific method for analyzing the severity of cracks in the roof slab is as follows: Roof slab images at various time points are read, and images of the crack initiation time and subsequent time points are extracted. These images are then renumbered chronologically as crack time point images, designated as 1, 2, ..., z, ..., s. Edge detection technology is used to extract the edge contours of each crack from each crack time point image, and the length of each crack in each crack time point image is obtained, denoted as L. zj Let j represent the number of the j-th crack, where j = 1, 2, ..., g. This is determined using the formula... The crack propagation rate ξ, L is obtained. (z-1)j Let Δt represent the length of the j-th crack in the image at the (z-1)-th crack time point, Δt represent the set time interval, s represent the number of crack time points, and g represent the number of cracks. Substitute these values ​​into the formula. The severity of the cracks in the top plate is obtained as θ, where ξ' represents the preset reference value for crack propagation rate. This helps to keep abreast of the crack development dynamics and take more targeted countermeasures to ensure operational safety.

[0046] S3. Impact analysis of roof cutting operation: The effect evaluation coefficient of the roof cutting operation is obtained by analyzing 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.

[0047] The specific method for detecting the pressure parameters of the coal seam working face after the roof cutting operation is as follows: First, select several equally spaced monitoring points on the coal seam working face, and use pressure sensors to detect the pressure at each monitoring point on the coal seam working face before the roof cutting operation, and record it as the pressure P at each monitoring point on the coal seam working face. f Let f represent the number of the f-th monitoring point, f = 1, 2, ..., k. The average pressure of the coal seam working face is obtained by averaging the pressure at each monitoring point, denoted as . Substitute it into the formula The pressure uniformity σ at the coal seam working face is obtained, where k represents the number of monitoring points. This helps to scientifically assess the impact of roof cutting operations on the pressure distribution at the coal seam working face, and provides data support for subsequent construction and safety assurance.

[0048] The second step involves performing roof cutting according to the predetermined plan. After the roof cutting is completed, several time points are selected at fixed intervals and recorded as monitoring time points. The pressure at each monitoring point on the coal seam face at each monitoring time point is measured and recorded as the pressure at each monitoring point on the coal seam face at each monitoring time point. The pressure uniformity at each monitoring time point on the coal seam face is analyzed using the method for analyzing the pressure uniformity of the coal seam face and recorded as σ'. x Let x represent the number of the x-th monitoring time point, x = 1, 2, ..., y. Adjacent monitoring time points are grouped together, and the pressure uniformity difference between adjacent monitoring time points in each group is calculated. These differences are then summed to obtain the pressure uniformity change value at the coal seam working face, denoted as Δσ'. Clearly defining the pressure uniformity change value allows for a better understanding of the pressure change trend and magnitude caused by roof cutting, providing a basis for optimizing roof cutting schemes and adjusting construction strategies, thereby improving construction efficiency and safety.

[0049] The specific analysis method for the effect evaluation coefficient of the coal seam roof cutting operation is as follows: The pressure uniformity σ at the coal seam working face and the pressure uniformity σ' at each monitoring time point of the coal seam working face are read respectively. x The variation value of pressure uniformity at the coal seam working face, Δσ', is substituted into the formula. The effect evaluation coefficient λ of the coal seam roof cutting operation is obtained, where η1 and η2 represent the preset pressure uniformity and pressure uniformity variation value weight factors, respectively, and Δσ'0 represents the preset pressure uniformity variation value reference value; it helps to optimize and improve the roof cutting operation, so as to improve the safety and efficiency of coal seam mining.

[0050] It should be noted that, in one 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, increase the risk of coal seam instability, and have an important impact on the roof cutting effect. The pressure uniformity change value mainly reflects the dynamic change of pressure uniformity over time or operation process. Although it can also reflect some trends and potential problems, the initial pressure uniformity state is more critical, so the pressure uniformity has a higher weight.

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

[0052] Please see Figure 2 As shown, the specific detection method for the roadway deformation and surrounding rock stability is as follows: First, when the roof collapses, video recording is performed on the roadway roof using a video recording device to obtain a video of the roadway roof. Several collection points are randomly selected on the roadway roof, and the positions of each collection point on the roadway roof in the first and last frames of the roadway roof video are marked to obtain the position points of each collection point on the roadway roof in the first and last frames. This helps to gain a deeper understanding of the specific details and dynamic changes of the roof collapse and to better grasp the behavioral characteristics of the roof.

[0053] The second step involves overlaying the last frame of the roadway roof video with the first frame of the roadway roof video, measuring the vertical distance between the last frame position point and the first frame position point of each acquisition point on the roadway roof, and recording this distance as the subsidence distance of each acquisition point on the roadway roof. This analysis yields the roadway deformation amount. This provides a quantitative indicator for assessing the stability and safety of the roadway, enabling the timely detection of potential risks, providing a scientific basis for roadway support design and maintenance, and improving the pertinence and effectiveness of support.

[0054] It should be noted that the specific analysis method for the roadway deformation is as follows: the subsidence distance of each sampling point on the roadway roof is read and denoted as d. c Let c represent the number of the c-th data collection point, where c = 1, 2, ..., 0. Substitute this number into the formula. The tunnel deformation ε is obtained, where d0 represents the reference subsidence distance of the preset collection point, and o represents the number of location points.

[0055] The third step involves simultaneously recording video of changes in the surrounding rock to obtain a video of the surrounding rock movement. Several locations on the surrounding rock are randomly selected, and the displacement distance of each location is obtained by analyzing the subsidence distance of each collection point on the roadway roof. The stability of the surrounding rock is thus analyzed. This helps to promptly identify unstable areas and potential hazards in the surrounding rock so that targeted reinforcement or preventive measures can be taken.

[0056] It should be noted that the specific method for analyzing the stability of the surrounding rock is as follows: read the displacement distance at each point on the surrounding rock, and denot it as d'. p Let p represent the number of the p-th location point, where p = 1, 2, ..., l. The average displacement distance of the surrounding rock is obtained by averaging the displacement distances of each location point, denoted as p. Substitute it into the formula The surrounding rock stability σ is obtained, where l represents the number of location points.

[0057] The specific analysis method for the feasibility of the aforementioned roof cutting and tunnel protection is as follows: read the tunnel deformation ε and the surrounding rock stability σ, and substitute them into the formula respectively. The feasibility χ of roof cutting and tunnel protection is obtained, where a1 and a2 represent the preset weighting factors of tunnel deformation and surrounding rock stability, respectively, and e represents the natural constant. This allows for a more accurate measurement of the feasibility of roof cutting and tunnel protection, reducing errors and uncertainties, making engineering planning more reasonable and reliable. At the same time, it allows for the prediction of potential risks and problems in advance, timely adjustment of strategies or the adoption of countermeasures, and reduction of engineering risks and costs.

[0058] It should be noted that, in one specific embodiment, a1 can be set to 0.5 and a2 can be set to 0.5. The roadway deformation directly reflects the actual changes in the roadway during and after the roof cutting process. If the deformation is too large, it will seriously affect the normal use and safety of the roadway, and directly affect the feasibility of the roof cutting and roadway protection scheme. The stability of the surrounding rock 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 not large, there may be significant problems later, which will have a profound impact on the long-term effect and feasibility of the roof cutting and roadway protection. Therefore, the roadway deformation and the surrounding rock stability have equal weights.

[0059] S5. Comprehensive Analysis of Support Pressure: Based on the bearing stability ξ of the coal pillar, the severity of roof cracks θ, the evaluation coefficient λ of the effect of coal seam roof cutting operation, and the feasibility χ of roof cutting and roadway protection, the influence evaluation index of support pressure on the coal seam fully mechanized longwall face is obtained. And provide feedback.

[0060] Please see Figure 3 As shown, the specific analysis method for the comprehensive analysis of the support pressure is as follows: The bearing stability ξ of the coal pillar, the severity of cracks θ in the roof, the evaluation coefficient λ of the effect of roof cutting operation, and the feasibility χ of roof cutting and roadway protection are respectively read and substituted into the formula. The evaluation index of the influence of support pressure on the fully mechanized coal seam mining face was obtained. w1, w2, w3, and w4 represent the bearing stability of the coal pillar, the severity of roof cracks, the evaluation coefficient of the effectiveness of roof cutting operations, and the weighting factor of the feasibility of roof cutting and roadway protection, respectively. The evaluation index of the impact of support pressure on the fully mechanized coal seam caving face is compared with the preset threshold of the support pressure impact evaluation index. If the evaluation index of the impact of support pressure on the fully mechanized coal seam caving face is greater than the preset threshold, it indicates that the support pressure may pose a significant risk to the working face. Conversely, it indicates that the impact of support pressure on the working face is relatively small, and the working face is in a relatively stable and safe state. Feedback on the impact of support pressure can clearly determine the degree of impact of support pressure on the working face and provide timely feedback. This helps to take targeted measures in advance to deal with potential risks, such as strengthening support and adjusting mining technology, thereby ensuring the safe and stable production of the working face.

[0061] It should be noted that, in one 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 the stability of the support structure of the entire working face and has a key impact on the distribution and magnitude of the support pressure. If the coal pillar is unstable, it may lead to abnormal changes in the support pressure, affecting the safety of the entire working face. The severity of roof cracks affects the integrity and bearing capacity of the roof, and thus affects the dispersion and transmission of support pressure. Severe cracks may lead to unreasonable distribution of support pressure, increasing potential risks. The effect of roof cutting operation will directly change the distribution of support pressure. If the roof cutting effect is good, it can effectively reduce problems such as support pressure concentration. The evaluation of the impact on support pressure is very important. Roof cutting and roadway protection mainly consider the impact on support pressure from a long-term and overall perspective. However, compared with the previous items, its direct impact may not be as urgent and critical. Therefore, the weights of the evaluation coefficients of the bearing stability of the coal pillar, the severity of roof cracks, and the effect of coal seam roof cutting operation are higher.

[0062] This invention collects and analyzes stress data of coal pillars to determine their bearing stability. It also analyzes crack propagation rates through image acquisition to assess the severity of roof cracks. Furthermore, it analyzes pressure parameters of the coal seam working face after roof cutting to obtain an evaluation coefficient for the effectiveness of the operation. Finally, it analyzes video recordings to determine roadway deformation and surrounding rock stability, thereby assessing the feasibility of roof cutting and roadway protection. This comprehensive analysis of the impact of fully mechanized coal seam caving face support pressure on the evaluation index allows for better implementation of targeted measures to ensure operational safety, optimize mining processes and procedures, and improve mining efficiency.

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

Claims

1. A method for simulating and analyzing the support pressure distribution in a fully mechanized longwall mining face of an extra-thick coal seam, characterized in that, Includes the following steps: S1. Bearing stability analysis: The stress data of the coal pillar is collected and analyzed to obtain the bearing stability of the coal pillar. The stress data of the coal pillar includes the proportion of high stress area at each time point and the stress value growth rate. S2. Crack Analysis: The crack initiation time is determined based on the degree of abnormality of the ultrasonic signal of the roof at each time point. Then, the crack propagation rate is analyzed by acquiring images to obtain the severity of the crack in the roof. S3. Impact analysis of roof cutting operation: Based on the pressure parameters of the coal seam working face after the roof cutting operation, the effect evaluation coefficient of the roof cutting operation is obtained. The pressure parameters include the pressure uniformity and the change value of the pressure uniformity. S4. Feasibility assessment of roof cutting and tunnel protection: When the roof collapses, the deformation of the tunnel and the stability of the surrounding rock are obtained by analyzing video recordings, and then the feasibility of roof cutting and tunnel protection is analyzed. S5. Comprehensive Analysis of Support Pressure: Based on the bearing stability ξ of the coal pillar, the severity of roof cracks θ, the evaluation coefficient λ of the effect of coal seam roof cutting operation, and the feasibility χ of roof cutting and roadway protection, the influence evaluation index of support pressure on the coal seam fully mechanized longwall face is obtained. And provide feedback.

2. The method for simulating and analyzing the support pressure distribution in a fully mechanized longwall mining face of an extra-thick coal seam according to claim 1, characterized in that: The specific method for detecting the stress data of the coal pillar is as follows: The first step involves selecting several detection points at different locations within the coal pillar according to a set interval, and simultaneously selecting several time points at a set time interval. Stress values ​​at each detection point within the coal pillar are collected using stress sensors and recorded as stress values ​​at each time point. These stress values ​​are then compared with a preset high-stress threshold. Detection points whose stress values ​​are greater than or equal to the preset high-stress threshold are selected and recorded as high-stress detection points. The boundaries of the high-stress regions at each time point are delineated by connecting these high-stress detection points, and the area of ​​the high-stress regions at each time point is obtained and denoted as M. i Let i represent the number of the i-th time point, i = 1, 2, ..., n, obtained through the formula... The proportion of high-stress regions at each time point of the coal pillar was obtained. M 煤柱 This represents the cross-sectional area of ​​the coal column; The second step involves averaging the stress values ​​at each detection point of the coal pillar at each time point to obtain the stress values ​​at each time point. Two adjacent time points are grouped together and denoted as each time point group. The stress values ​​of each time point group are subtracted to obtain the stress value difference of each time point group. The stress value growth rate of each time point group is obtained by dividing the stress value difference of each time point group by a set time interval. The average value is then used to calculate the stress value growth rate of the coal pillar, denoted as V.

3. The method for simulating and analyzing the support pressure distribution in a fully mechanized longwall mining face of an extra-thick coal seam according to claim 2, characterized in that: The specific analysis method for the load-bearing stability analysis is as follows: The proportion of high-stress areas at each time point of the coal pillar was read separately. The rate of increase V of stress in the coal pillar is substituted into the formula. The bearing stability ξ of the coal pillar is obtained, V' represents the preset reference value of stress value growth rate, φ1 and φ2 represent the preset proportion of high stress area and the weight factor of stress value growth rate, respectively, and n represents the number of time points.

4. The method for simulating and analyzing the support pressure distribution in a fully mechanized longwall mining face of an extra-thick coal seam according to claim 3, characterized in that: The specific analytical method for determining the crack initiation time point is as follows: The first step involves acquiring images of the roof slab at various time points using a camera. Simultaneously, several detection points are selected on the roof slab at predetermined distances and designated as "roof slab detection points." Coupling agent is applied to each detection point. An ultrasonic flaw detector then transmits ultrasonic signals of fixed intensity to each detection point at each time point. A probe receives the reflected ultrasonic signals from the roof slab, and the duration of the ultrasonic signal transmission and reception at each detection point at each time point is recorded as T. im Let m represent the number of the m-th roof detection point, m = 1, 2, ..., q. The intensity of the ultrasonic signal reflected from each roof detection point at each time point is denoted as G. im Through formula The degree of anomaly δ of the ultrasonic signal of the roof at each time point was obtained. i G represents a fixed-intensity ultrasonic signal, and q represents the number of detection points on the top plate. These represent the preset duration of the ultrasonic signal from transmission to reception and the weighting factor of the ultrasonic signal intensity, respectively. The second step is to compare the degree of abnormality of the ultrasonic signal of the roof at each time point with the preset threshold for the degree of abnormality of the ultrasonic signal. If the degree of abnormality of the ultrasonic signal of the roof at a certain time point is greater than or equal to the preset threshold for the degree of abnormality of the ultrasonic signal, it is determined that there is a crack in the roof at that time point. The time point when the degree of abnormality of the ultrasonic signal of the roof is greater than or equal to the preset threshold for the degree of abnormality of the ultrasonic signal is recorded as the crack start time point.

5. The method for simulating and analyzing the support pressure distribution in a fully mechanized longwall mining face of an extra-thick coal seam according to claim 4, characterized in that: The specific method for analyzing the severity of the cracks in the top slab is as follows: Read the top plate images at various time points, extract the top plate images at the crack initiation time point and subsequent time points, and renumber them chronologically as crack time point images, numbered 1, 2, ..., z, ..., s. Use edge detection technology to extract the edge contours of each crack from each crack time point image, and obtain the length of each crack in each crack time point image, denoted as L. zj Let j represent the number of the j-th crack, where j = 1, 2, ... g., determined by the formula... The crack propagation rate ξ, L is obtained. (z-1)j Let Δt represent the length of the j-th crack in the image at the (z-1)-th crack time point, Δt represent the set time interval, s represent the number of crack time points, and g represent the number of cracks. Substitute these values ​​into the formula. The severity of the cracks in the top plate is obtained as θ, where ξ' represents the preset reference value for crack propagation rate.

6. The method for simulating and analyzing the support pressure distribution in a fully mechanized longwall mining face of an extra-thick coal seam according to claim 1, characterized in that: The specific method for detecting the pressure parameters of the coal seam working face after the top-cutting operation is as follows: The first step is to select several equally spaced monitoring points at the coal seam working face, and use pressure sensors to detect the pressure at each monitoring point before the roof cutting operation, which is recorded as the pressure P at each monitoring point of the coal seam working face. f Let f represent the number of the f-th monitoring point, f = 1, 2, ..., k. The average pressure of the coal seam working face is obtained by averaging the pressure at each monitoring point, denoted as . Substitute it into the formula The pressure uniformity σ at the coal seam working face is obtained, where k represents the number of monitoring points; The second step involves performing roof cutting according to the predetermined plan. After the roof cutting is completed, several time points are selected at fixed intervals and recorded as monitoring time points. The pressure at each monitoring point on the coal seam face at each monitoring time point is measured and recorded as the pressure at each monitoring point on the coal seam face at each monitoring time point. The pressure uniformity at each monitoring time point on the coal seam face is analyzed using the method for analyzing the pressure uniformity of the coal seam face and recorded as σ'. x x represents the number of the xth monitoring time point, x = 1, 2, ..., y. At the same time, adjacent monitoring time points are grouped together. The pressure uniformity difference between adjacent monitoring time points in each group is obtained by subtracting the pressure uniformity of adjacent monitoring time points in each group. The sum of these values ​​is used to obtain the pressure uniformity change value of the coal seam working face, which is denoted as Δσ'.

7. The method for simulating and analyzing the support pressure distribution in a fully mechanized longwall mining face of an extra-thick coal seam according to claim 6, characterized in that: The specific analysis method for the effectiveness evaluation coefficient of the coal seam roof cutting operation is as follows: The pressure uniformity σ at the coal seam working face and the pressure uniformity σ' at each monitoring time point of the coal seam working face were read respectively. x The variation value of pressure uniformity at the coal seam working face, Δσ', is substituted into the formula. The effect evaluation coefficient λ of the coal seam roof cutting operation is obtained, where η1 and η2 represent the preset pressure uniformity and pressure uniformity variation value weight factors, respectively, and Δσ'0 represents the preset pressure uniformity variation value reference value.

8. The method for simulating and analyzing the support pressure distribution in a fully mechanized longwall mining face of an extra-thick coal seam according to claim 1, characterized in that: The specific methods for detecting the tunnel deformation and surrounding rock stability are as follows: The first step is to record the tunnel roof with video recording equipment when the roof collapses, and then select several collection points on the tunnel roof to mark the position of each collection point in the first and last frames of the tunnel roof video. The second step is to overlay the last frame of the tunnel roof video with the first frame of the tunnel roof video, measure the vertical distance between the last frame position point and the first frame position point of each collection point of the tunnel roof, record it as the subsidence distance of each collection point of the tunnel roof, and analyze the tunnel deformation. The third step involves simultaneously recording video of the changes in the surrounding rock to obtain a video of the surrounding rock movement. Several locations on the surrounding rock are randomly selected, and the displacement distance of each location is obtained by analyzing the subsidence distance of each collection point on the roadway roof. The stability of the surrounding rock is then analyzed.

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

10. The method for simulating and analyzing the support pressure distribution in a fully mechanized longwall mining face of an extra-thick coal seam according to claim 1, characterized in that: The specific analysis method for the comprehensive analysis of the support pressure is as follows: The bearing stability ξ of the coal pillar, the severity of cracks in the roof θ, the evaluation coefficient of the effect of roof cutting operation λ, and the feasibility of roof cutting and roadway protection χ are respectively read and substituted into the formula. The evaluation index of the influence of support pressure on the fully mechanized coal seam mining face was obtained. w1, w2, w3, and w4 represent the bearing stability of the coal pillar, the severity of cracks in the roof, the evaluation coefficient of the effect of coal seam roof cutting operation, and the weighting factor of the feasibility of roof cutting and roadway protection, respectively. The evaluation index of the impact of the support pressure on the coal seam fully mechanized caving face is compared with the preset threshold of the support pressure impact evaluation index. If the evaluation index of the impact of the support pressure on the coal seam fully mechanized caving face is greater than the preset threshold of the support pressure impact evaluation index, it indicates that the support pressure may pose a greater risk to the working face. Conversely, it indicates that the impact of the support pressure on the working face is relatively small, and the working face is in a relatively stable and safe state. Feedback is provided on the impact of the support pressure.

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