A method for evaluating overall water-blocking performance of floor based on permeability evolution

Abstract

A kind of floor overall water-blocking performance evaluation method based on permeability evolution, obtain the hydrogeological conditions of the study area and establish a complete database;Combined with geological and mining conditions, a composite floor plastic slip failure mechanics model is established, combined with multiple monitoring techniques field verification, determine the maximum development depth of mining damage zone;Using numerical simulation to determine the distribution of mining damage zone, combined with a variety of test methods to verify, accurately define its range;Through numerical simulation to determine the mining stress path and equivalent transformation into laboratory loading path, carry out seepage experiment, establish the mathematical model of permeability evolution;Combined with the calculation of key parameters, the pressure guide lifting zone range is determined, and the actual thickness is determined by field water level observation calibration;Using the weighted average method to calculate the equivalent water-blocking thickness, combined with the hydrogeological conditions to establish three-level evaluation standard, realize the quantitative evaluation and risk classification of water-blocking performance. This method can realize the accurate quantitative evaluation of floor overall water-blocking performance evaluation.

Description

Technical Field

[0001] This invention belongs to the field of intelligent mine safety mining technology, specifically involving a method for evaluating the overall water-blocking performance of the bottom plate based on permeability evolution. Background Technology

[0002] As the core support of my country's energy security, coal mining faces an increasingly severe threat from high-pressure water as the depth of mines continues to extend at a rate of 8-12 meters per year. Water inrush from the floor has become a key problem restricting the safe and efficient mining of coal resources, directly affecting the efficiency of coal resource development and the safety of mine production.

[0003] Ordovician limestone aquifers are characterized by high pressure (5-8 MPa), strong water abundance, and stable recharge. Furthermore, the close proximity of these aquifers to coal seams means that approximately 60% of mines in my country are directly threatened by confined water in the floor. Historically, numerous major floor water inrush accidents have occurred, rendering about 40% of coal resources unusable and creating substantial amounts of idle resources, resulting in severe resource waste and safety hazards. The water-blocking performance of the floor aquitard is the core barrier against confined water inrush into mines. However, during coal seam mining, the floor rock mass undergoes a complex stress evolution path of stress concentration, decompression, and recompaction. This process easily triggers the initiation, expansion, and connection of internal fissures, leading to drastic changes in rock permeability. The originally intact and continuous aquitard may gradually lose its water-blocking function. This dynamic evolution of the aquitard's water-blocking performance induced by mining directly determines the overall water-blocking capacity of the floor and is a core research object for floor water inrush risk prevention and control.

[0004] Currently, existing methods for evaluating the water-blocking performance of the floor slab still have significant limitations and shortcomings, making it difficult to meet the safety requirements of mining under the threat of deep, high-pressure water. These limitations are mainly reflected in three aspects: First, the evaluation methods are mostly based on qualitative judgments using static geological parameters, failing to fully consider the dynamic coupling relationship between mining stress and rock permeability evolution, and ignoring the heterogeneous differences in permeability after mining. This leads to significant deviations between the evaluation results and actual mining conditions, making it difficult to accurately reflect the true water-blocking state. Second, the detection methods for core evaluation parameters such as floor slab damage depth and the extent of mining-induced damage zones are relatively simple, lacking a multi-faceted and multi-dimensional collaborative verification mechanism, resulting in insufficient parameter detection accuracy and an inability to provide reliable data support for water-blocking performance evaluation. Third, the calculation of equivalent water-blocking thickness lacks a scientifically sound weight allocation system and fails to incorporate graded evaluation based on differences in mine geological structures, making it difficult to achieve accurate grading and targeted treatment of floor slab water-blocking risks, and failing to meet the prevention and control needs under different geological conditions.

[0005] Given the complex geological conditions of composite floor strata, existing technologies are insufficient to accurately depict the dynamic evolution of the water-blocking properties of aquifers during mining operations. This hinders the provision of scientific and reliable evaluation criteria for water-blocking performance in pressurized coal seams, severely restricting the safe and efficient development of coal resources and the release of stagnant resources in areas threatened by high-pressure water. Therefore, there is an urgent need for a dynamic and quantitative method for evaluating the overall water-blocking performance of the floor strata based on permeability evolution. Summary of the Invention

[0006] To address the problems existing in the prior art, this invention provides a method for evaluating the overall water-blocking performance of a base plate based on permeability evolution. This method can achieve accurate quantitative evaluation of the overall water-blocking performance of the base plate, and is particularly suitable for quantitative evaluation of the water-blocking performance of composite base plates.

[0007] To achieve the above objectives, this invention provides a method for evaluating the overall water-blocking performance of a substrate based on permeability evolution, comprising the following steps: Step 1: Through geophysical exploration, borehole exploration, and the summarization and organization of existing geological data, systematically obtain the hydrogeological conditions of the study area and establish a complete hydrogeological information database; Step 2: Based on geological and mining conditions, establish a composite floor plastic slip failure mechanical model under coal seam mining conditions to clarify the floor plastic slip failure depth and state corresponding to different floor types, different failure states, and mining conditions; at the same time, conduct multi-dimensional field verification by combining borehole detection method, network parallel electrical resistivity tomography monitoring technology, and distributed optical fiber monitoring technology to determine the maximum development depth of the floor mining failure zone; Step 3: Use numerical simulation to determine the distribution range of the mining-induced damage zone of the floor, and combine borehole water injection test, borehole inspection and transient electromagnetic method to conduct full-space multi-parameter collaborative verification to accurately define the distribution range of the mining-induced damage zone of the floor. Step 4: Combine numerical simulation to determine the mining stress path of the rock strata in the mining-induced damage zone, and convert the mining stress path into an equivalent laboratory loading path. Conduct seepage experiments on each rock stratum within the mining-induced damage zone under the mining stress path to clarify the permeability evolution characteristics of rock samples at different stress stages and establish mathematical models for the permeability evolution of rock samples at different stress stages. Step 5: Calculate the development range of the pressure-bearing guide zone of the bottom plate by combining key parameters, and calibrate it by combining on-site water level observation data to comprehensively determine the actual thickness of the pressure-bearing guide zone of the bottom plate within the mining area; Step 6: Based on the actual thickness and permeability evolution characteristics of each rock layer, the weighted average method is used to calculate the equivalent water-blocking thickness of the base plate; combined with the presence of tectonic structures and other hydrogeological conditions of the base plate, a three-level water-blocking performance evaluation standard is established to achieve quantitative evaluation of the overall water-blocking performance of the base plate and safety risk classification.

[0008] Furthermore, in order to provide systematic and comprehensive basic data support for subsequent evaluation work, in step 1, the regional hydrogeological conditions include the distance between the coal seam and the aquifer, the lithological composition and thickness of the aquitard, the rock strength and initial fracture development characteristics, the distribution and scale of geological structures, and the water-rich state and water pressure of the aquifer. To provide accurate and targeted detection data support for evaluating the water-blocking performance of the foundation, the geophysical exploration employs a combined approach of transient electromagnetic method and high-density resistivity method, focusing on detecting the distribution characteristics of adverse geological bodies. The borehole exploration focuses on obtaining the distance between the coal seam and the aquifer, the lithological composition and thickness of the aquitard, and the rock layer strength and initial fracture development characteristics. Simultaneously, the hydrogeological parameters in the detection results meet the following measurement accuracy requirements: aquifer water pressure error ≤ 0.05 MPa, rock layer thickness measurement error ≤ 5%, and rock layer strength test error ≤ 3%. To accurately output data related to the distribution of mining-induced damage zones in the floor and provide reliable support for subsequent evaluation, in step 3, the numerical simulation uses FLAC3D software to establish a fluid-structure interaction model. The model dimensions are determined based on the working face strike, dip length, and floor detection depth, with a length × width × height of no less than 400m × 400m × 200m. The fluid-structure interaction model incorporates a stress-damage-seepage interaction model to characterize the evolution of rock mass permeability characteristics. A double-yield model is used to simulate the compaction characteristics of collapsed rock blocks in the goaf. The simulation step size is set to 5-10m. Displacement constraints are applied around the model and at the bottom, an equivalent overlying stratum pressure is applied at the top, and actual water pressure is applied to the floor aquifer. The model calculation convergence criterion is an unbalanced force ratio ≤ 1 × 10⁻⁶. -5 Output the longitudinal distribution range and lateral boundary of the mining-induced damage zone; In order to accurately calibrate the development range of the pressure-bearing guide zone of the bottom plate and ensure the reliability of the relevant parameter definition, in step 5, the calibration of the on-site water level observation data is achieved by arranging at least 3 calibration boreholes. The depth of the calibration boreholes is not less than 1.5 times the calculated guide height, and the bottom of the boreholes is located 5-10m above the top plate of the aquifer. The water level elevation in the boreholes is measured by water level measuring ropes or pressure sensors with a measurement accuracy of ≤0.1m. The water level is continuously observed for 3 hydrological cycles, and the actual guide height is calculated by taking the stable water level elevation.

[0009] Furthermore, in order to accurately quantify the plastic slip failure depth of the base plate through a dedicated formula, with clear parameter definitions and a design that fits reality, and to support the accurate calculation of subsequent failure depth, in step 2, the composite base plate plastic slip failure mechanical model includes three types of base plates: single-structure type, double-layer structure type, and soft-hard composite type. Among them, the double-layer structure type base plate includes three failure states, and the soft-hard composite type base plate includes five failure states. The model calculation parameters are calibrated through field geological survey data and laboratory rock sample test data. The plastic slip failure depth of the base plate is obtained according to formula (1). ; (1); In the formula, This represents the total thickness of the aquitard from the bottom of the coal seam to the top of the aquifer, in meters (m). , , , , The fitting parameters are for 5 different models; , , These are the internal friction angles of the bottom rock strata for single-striated, double-layered, and soft-hard composite types, respectively, in degrees. The crack width is expressed in mm. For crack roughness; This refers to the hydrostatic pressure of the water-containing material, expressed in MPa. Initial permeability of the rock formation, in meters. 2 ; Permeability of the rock strata after mining, in meters. 2 ; The rock mass compressive strength is expressed in MPa. This represents the maximum principal stress actually borne by the rock strata, expressed in MPa. , These represent the thicknesses of each rock layer in the double-layer structure and the soft-hard composite base plate, respectively, all in meters (m).

[0010] Furthermore, in order to accurately determine the maximum development depth of the mining-induced damage zone in the base plate, the following process is used in step 2 to conduct on-site verification of the dimensionality by combining borehole detection method, network parallel electrical resistivity tomography (EPT) monitoring technology, and distributed optical fiber monitoring technology: S21: Drilling detection method: By arranging at least 3 monitoring boreholes, the borehole depth is not less than 1.2 times the predicted failure depth, and the borehole verticality error is ≤1°, the borehole inspection instrument is used to observe the crack penetration status at different depths in the base plate. S22: Network parallel electrical resistivity monitoring method: Determine the fracture development depth by the area of ​​abnormally high apparent resistivity. In the mining roadway, at least 4 measuring lines are arranged using a combination of borehole and roadway methods. Each measuring line is equipped with more than 16 electrodes. The electrode spacing is ≤4m, the apparent resistivity detection accuracy is ≤10Ω・m, and the data acquisition interval is ≤4h. S23: Distributed fiber optic monitoring technology: Fiber optic sensors are deployed vertically along the base plate with a sensor spacing of ≤5m, strain measurement accuracy of ≤10με, and data sampling frequency of ≥1Hz. Stress concentration depth is identified by strain abrupt change characteristics. S24: Determine the maximum development depth of the mining-induced damage zone of the base plate by combining data from three types of monitoring.

[0011] Furthermore, in order to accurately define the distribution range of the base plate mining damage zone, the process of multi-parameter collaborative verification in the entire space in step 3 is as follows: S31: Borehole water injection test method: Conduct water injection tests in sections in at least two monitoring boreholes in the key area. The test section is 2-3m long. The constant pressure method is used to control the water injection pressure at 0.5-1.0MPa. Record the leakage of rock mass per unit length. When the leakage is ≥5L / (min・m), it is determined to be a section with developed water-conducting fractures. S32: Drilling inspection method: Using a drilling inspection instrument with a resolution of ≥0.1mm, continuous observation is carried out along the entire depth of the monitored borehole to record the fracture aperture, density and continuity status, and to divide the section into fracture-free section, micro-fracture section and densely fractured section. S33: Transient electromagnetic method: Employs a multi-turn small loop transmitter with a transmission frequency of 50-500Hz, a detection depth of not less than 1.2 times the predicted damage zone depth, and detection angles covering 0°, 45°, 90°, and 135°; data processing uses a regularized inversion algorithm with an inversion accuracy of ≤10%, and the range of rock mass property changes is determined by the apparent resistivity difference. S34: By combining the overlapping areas of the numerical simulation results and the three types of verification data, the top boundary, bottom boundary and lateral extension range of the mining-induced damage zone of the base plate are determined, so as to accurately define the distribution range of the mining-induced damage zone. The rules for defining the distribution range of the mining-induced damage zone in the floor are as follows: the top boundary is the bottom boundary of the mining-induced damage zone; the bottom boundary is the depth that meets any of the following conditions: the depth where the rock mass damage variable in the numerical simulation is ≤0.1, the depth where the leakage in the borehole water injection test is <1L / (min・m), the depth where there are no obvious cracks when the borehole is inspected, and the depth where the transient electromagnetic apparent resistivity difference is <50Ω・m; the lateral extension range is based on the working face boundary and extends to both sides for no less than 20m, or to the stress stability zone of the coal pillar.

[0012] Furthermore, in order to obtain a complete four-stage stress path and clarify the key parameters of each stage, so as to provide accurate and comprehensive data support for the subsequent equivalent transformation of stress path, in step 4, the mining stress path is obtained by the following numerical simulation method: monitoring points are set up in layers according to rock strata within the mining damage zone, with at least 3 monitoring points set up in each rock stratum and a spacing of 5m. The stress evolution data of the monitoring points during the working face advance process is extracted to obtain a complete four-stage stress path including the original rock stress stage, the advanced stress concentration stage, the mining pressure relief stage, and the stress environment adjustment stage, and to clarify the stress magnitude, stress change rate and duration of each stage; To accurately match the stress magnitude and equivalent amplified stress change rate between the field and the laboratory, and to ensure that the mechanical response of the laboratory loading path is consistent with that of the field mining stress path, thus providing a realistic loading basis for subsequent seepage experiments, the equivalent transformation follows the principles of stress evolution equivalence and time scale compression. The stress magnitude at each stage is kept consistent with the ratio coefficient between the field stress and the strength of the laboratory rock sample, and the stress change rate is equivalently amplified according to the compression ratio of the field stress evolution cycle and the laboratory test cycle, ensuring the consistency of the mechanical response between the laboratory loading path and the field mining stress path. To ensure accurate and reliable experimental results that closely reflect actual field conditions, the seepage experiment employed a ROCKTOP multi-field coupling test system. This system meets the following requirements: maximum axial load range ≥2000kN, confining pressure control accuracy ≤±0.1MPa, and seepage flow rate measurement accuracy ≤1×10⁻⁶. -8 m 3 / s, temperature control accuracy ≤±1℃, the temperature is kept constant during the test to eliminate the influence of temperature on permeability; the rock samples are prepared by core sampling from each rock layer in the mining-induced damage zone, the sample size is Φ50mm×100mm, and 3 parallel samples are prepared for each group of rock layers to reduce dispersion; the confining pressure of the test is set according to the confining pressure value of the corresponding rock layer in the field, the permeation medium is deionized water, the permeation pressure is controlled at 0.5-5MPa, the parameters are recorded in real time during the loading process, and the loading-holding-depressurization-reloading is completed in sequence according to the four stages of the mining-induced stress path; To provide solid data support for the subsequent establishment of the permeability evolution model, the permeability evolution characteristics are determined by calculating the permeability values ​​at each stress stage. The permeability calculation employs either the steady-state method or the transient pressure pulse method, where the permeability is ≤1×10⁻⁶. -17 m 2 The transient method was used; the steady-state method met the flow stability criterion (flow deviation ≤ 5% for 3 consecutive measurements); the transient pressure pulse method required recording the pressure decay curves of upstream and downstream over time, and calculating the permeability by fitting the decay curves; the permeability value at each stress stage was taken as the arithmetic mean of 3 parallel samples, with a relative standard deviation ≤ 15%; To accurately characterize the evolution of permeability at different stress stages and enhance the reliability and applicability of the model, the permeability evolution mathematical model uses stress parameters as independent variables and permeability as the dependent variable, employing exponential, power, or linear functions for fitting. The model's goodness of fit R0 is [value missing]. 2 ≥0.85, ensuring accurate characterization of the evolution of permeability at different stress stages.

[0013] Furthermore, to ensure the accuracy and reliability of the acquired key parameters and to provide solid data support for subsequent calculations, in step 5, the key parameters are obtained through a combination of on-site testing and laboratory analysis, as follows: S51: Aquifer water pressure is monitored by hydrological observation wells. At least two representative wells are selected and monitored continuously for more than 72 hours. The stable water pressure value is used as the basis for calculation. The monitoring accuracy is ≤0.05MPa. S52: The width of the original fractures inside the rock strata is observed by stereomicroscope after core drilling. No less than 3 sections are observed for each rock core, and no less than 10 fractures are measured in each section. The arithmetic mean is taken as the fracture width of the rock strata, and the measurement accuracy is ≤0.01mm. S53: The roughness of the original fractures inside the rock strata is obtained by scanning the fracture surface profile. A laser scanner is used to scan the fracture surface with a scanning resolution of ≤50μm. The roughness is characterized by the undulation coefficient of the profile curve, with the undulation coefficient ranging from 0.1 to 5.0μm.

[0014] Furthermore, in order to accurately fit the permeability evolution law under axial stress and confining pressure, in step 4, a mathematical model of permeability evolution is established according to formula (2); (2); In the formula, k is the permeability, α and β are two different fitting parameters, σ1 is the axial stress in MPa, and σ3 is the confining pressure in MPa. In order to quickly and accurately solve the core parameters and provide a clear and reliable calculation basis for determining the actual thickness of the pressure-bearing guide strip in the subsequent step, the development range of the pressure-bearing guide strip of the bottom plate is calculated according to formula (3) in step 5. ; (3).

[0015] Furthermore, in order to accurately obtain the equivalent water-blocking thickness of the waterproof layer, the calculation process for the equivalent water-blocking thickness of the base plate in step 6 is as follows: S161: Determine basic parameters; execute steps 1 to 5 to determine the thickness of the mining-induced failure zone, mining-induced damage zone, original rock stress zone, and pressure-bearing conduction zone, as well as the evolution law of rock permeability in the mining-induced damage zone. S162: Quantification of mining-induced water resistance coefficient; introduction of mining-induced water resistance coefficient The water-blocking coefficient is used to quantify the change in the water-blocking characteristics of the rock strata under mining action. According to formula (4), the water-blocking coefficient is calculated by the ratio of the initial permeability of the rock strata to the permeability at a certain moment. : (4); In the formula, k0 is the initial permeability of the rock layer; k is the permeability of the rock layer.

[0016] S163: General Calculation of Equivalent Water-Blocking Thickness; The equivalent water-blocking thickness is used to characterize the total thickness of the rock layer with a mining-induced water-blocking coefficient of 1. For rock layers with reduced water-blocking capacity, the equivalent thickness is the product of the rock layer thickness and the mining-induced water-blocking coefficient. Based on this, a general calculation is performed according to formula (5) to obtain the equivalent water-blocking thickness of the aquitard. ; (5) In the formula, The number of rock strata between the bottom of the coal seam and the top of the aquifer; The water resistance coefficient for mining of each rock stratum; The thickness of each rock layer is given in meters (m). S164: Specific calculations based on the lower four-zone theory; combining the lower four-zone theory, all rock strata from the bottom of the coal seam to the top of the aquifer are divided into mining-induced failure zone, mining-induced damage zone, original rock stress zone, and pressure-bearing uplift zone, with corresponding thicknesses as follows: , , , Based on the water-blocking characteristics of each stratum, determine its equivalent water-blocking contribution coefficient: the water-blocking capacity of the mining-induced failure zone and the confined uplift zone is extremely low or completely lost, so set its mining-induced water-blocking coefficient. The mining-induced damage zone, although not destroyed by coal seam mining disturbance, has gradually developed internal fractures, resulting in varying degrees of decrease in water-blocking capacity. A mining-induced water-blocking coefficient is set based on the degree of decrease in water-blocking capacity. Located between 0 and 1; the original rock stress zone has not been disturbed by mining, and its water-blocking capacity has not changed. Assume its mining-induced water-blocking coefficient is... Pressure-bearing riser zone It does not have an effective water-blocking function, so let its mining-induced water-blocking coefficient be... Based on this, the equivalent water-blocking thickness of the waterproof layer is obtained by performing specific calculations according to formula (6). ; (6).

[0017] Furthermore, in order to achieve accurate quantitative evaluation of water-blocking performance and precise risk classification, the process of quantitative evaluation and safety risk classification of the overall water-blocking performance of the base plate in step 6 is as follows: S261: Calculation of the bottom plate water inrush coefficient; Considering the characteristics of the water-resistant layer, the corrected bottom plate water inrush coefficient is calculated based on the equivalent water-resistant thickness according to formula (7). ; (7); In the formula, The aquifer pressure is expressed in MPa. S262: Water-blocking performance evaluation; Based on on-site engineering practice and water-blocking layer protection requirements, a three-level water-blocking performance evaluation standard is adopted, with the water inrush coefficient corresponding to the equivalent water-blocking thickness as the core indicator: water inrush coefficient ≤ 0.04MPa / m is excellent, corresponding to low risk, and normal mining is possible; 0.04MPa / m < water inrush coefficient ≤ 0.06MPa / m is moderate, corresponding to medium risk, and enhanced monitoring measures are required; water inrush coefficient > 0.6MPa / m is poor, corresponding to high risk, and remediation measures are required before mining; the thickness of the key water-blocking layer must meet the basic requirement of not less than 20m. If it is less than 20m, further enhanced remediation measures are required.

[0018] This invention provides a method for evaluating the overall water-blocking performance of a substrate based on permeability evolution. Through systematic process design and technological innovation, it has the following advantages: 1. Solid foundational data and accurate and reliable definition of key parameters: Through geophysical exploration, borehole exploration, and the summarization of existing geological data, the hydrogeological conditions of the study area are systematically obtained, and a complete hydrogeological information database is established, providing a reliable data foundation for subsequent accurate evaluation work and avoiding evaluation bias caused by fragmented basic information. Based on this, a collaborative verification system integrating mechanical models, numerical simulations, and multi-site testing was constructed for the core parameters during the mining process of the floor slab, achieving precise parameter definition: First, a composite floor slab plastic slip failure mechanical model was established under coal seam mining conditions, combining geological and mining conditions, clarifying the depth and state of floor slab plastic slip failure under different working conditions. Simultaneously, multi-dimensional field verification was conducted using borehole detection, network parallel electrical resistivity tomography (EPM) monitoring, and distributed fiber optic monitoring to determine the maximum development depth of the floor slab mining-induced damage zone. Second, numerical simulation methods were used to clarify the distribution range of the floor slab mining-induced damage zone. Combined with borehole water injection tests, borehole inspection, and transient electromagnetic methods, multi-parameter collaborative verification was conducted throughout the space to further accurately define the distribution of the mining-induced damage zone. Third, the development range of the floor slab pressure-bearing and lifting zone was calculated based on key parameters, and its actual thickness was determined through calibration using field water level observation data. The application of these multi-technical cross-verification methods effectively avoided the blind spots and errors of single detection methods, significantly improving the accuracy of key parameter definition and providing reliable technical support for subsequent evaluation work.

[0019] 2. An innovative stress path equivalence method was developed, accurately capturing the evolution law of permeability. This invention breaks through the limitations of traditional static evaluation, focusing on the dynamic coupling relationship between mining stress and the permeability of the base strata. It innovatively adopts a technical approach combining in-situ stress path equivalence transformation with laboratory seepage experiments: First, the actual mining stress path of the mining-damaged strata is determined through numerical simulation. Then, this in-situ mining stress path is equivalently transformed into a laboratory loading path. Seepage experiments are then conducted on each stratum within the mining-damaged zone under the mining stress path, accurately capturing the permeability evolution characteristics of rock samples at different stress stages and establishing corresponding mathematical models of permeability evolution. This process achieves precise alignment between in-situ mining conditions and laboratory experiments, overcoming the shortcomings of traditional experiments that are detached from the actual mining stress environment and lack accurate characterization of permeability evolution laws. It can truly reflect the dynamic changes in the water-blocking performance of the base strata during mining, providing core theoretical and experimental support for subsequent quantitative evaluation.

[0020] 3. The calculation process for equivalent water-blocking thickness is scientific, and the evaluation criteria are adapted to complex working conditions. Based on the permeability evolution model, the innovative method adopts a weighted average method based on the ratio of initial permeability to post-mining permeability, combined with the actual thickness and permeability evolution characteristics of each rock layer, to calculate the equivalent water-blocking thickness of the base plate. This calculation method emphasizes the water-blocking contribution weight of low-permeability rock layers, effectively overcoming the problems of traditional weighted average methods neglecting the differences in water-blocking capacity of different rock layers and the disconnect between the calculated results and the actual water-blocking effect. This ensures that the equivalent water-blocking thickness can truly reflect the overall water-blocking level of the base plate, significantly improving the scientific nature of the calculation. Simultaneously, considering specific hydrogeological conditions such as the presence of structural features in the base plate, a three-level water-blocking performance evaluation standard is established, achieving precise alignment between quantitative evaluation of the base plate's water-blocking performance and safety risk classification, providing a clear basis for developing targeted prevention and control measures for areas with different risk levels.

[0021] 4. Breaking through the limitations of traditional evaluation methods, this invention demonstrates significant engineering applicability and promotional value. It overcomes the limitations of existing technologies that rely on qualitative evaluation based on static geological parameters. Instead, it focuses on the dynamic evolution of the floor's water-blocking performance during mining operations. Through quantitative analysis, precise verification, and scientific calculation throughout the entire process, it achieves dynamic quantitative evaluation of the overall water-blocking performance of the floor under mining conditions and precise classification of safety risks. The evaluation results are more closely aligned with actual mining conditions, providing a scientific and reliable technical basis for safe coal seam mining under the threat of high-pressure water, effectively ensuring safe and efficient mine operation. Meanwhile, the method is highly operable and widely applicable: it has designed suitable technical processes for different floor types such as single structure, double-layer structure, and soft-hard composite type, and quantified the operation parameters and criteria of each step. It does not require complex special equipment and is easy to promote and apply in the field. It is especially suitable for the prediction of floor water hazard risk in deep coal seam mining under the threat of high-pressure water. It is highly targeted for the quantitative evaluation of the water-blocking performance of composite floor. It has important engineering significance and application value for liberating stagnant coal resources under the threat of high-pressure water and reducing the risk of floor water hazard prevention and control, and has broad prospects for promotion.

[0022] 5. The core value is highlighted, contributing to improved quality and efficiency in water hazard prevention and control. This invention ultimately achieves precise quantitative evaluation of the water-blocking performance of composite foundations threatened by confined water. By refining key parameters, evolution patterns, and differentiating evaluation standards, it effectively addresses the pain points of existing technologies, such as difficulty in adapting to complex geological conditions of composite foundations and insufficient reliability of evaluation results. Its established three-level evaluation standard can clearly define the water hazard risk level of different areas, providing support for the development of targeted treatment plans such as "normal mining, enhanced monitoring, and grouting reinforcement" on-site. This avoids the blind application of water hazard prevention and control measures, improving the effectiveness of water hazard prevention and control while reducing costs, achieving a dual improvement in safe mining and economic benefits, further highlighting the engineering practical value and promotional significance of the method.

[0023] This method enables precise quantitative evaluation of the overall water-blocking performance of the coal seam floor, and is particularly suitable for quantitative evaluation of the water-blocking performance of composite coal seams. It effectively addresses the problem that existing technologies struggle to accurately characterize the evolution of water-blocking characteristics of aquifers during mining, especially in complex geological conditions with composite coal seams, under the threat of high-pressure water. This makes it difficult to provide a scientific basis for evaluating pressurized mining. This invention has significant engineering practical value and broad application value for ensuring safe and efficient coal mining under the threat of high-pressure water, promoting the release of stagnant coal resources, and improving mine safety production levels. Attached Figure Description

[0024] Figure 1 This is a schematic diagram of the process of the present invention. Detailed Implementation

[0025] The present invention will be further described below with reference to the embodiments.

[0026] like Figure 1 As shown, this invention provides a method for evaluating the overall water-blocking performance of a substrate based on permeability evolution, comprising the following steps: Step 1: Through geophysical exploration, borehole exploration, and the summarization and organization of existing geological data, systematically obtain the hydrogeological conditions of the study area and establish a complete hydrogeological information database; In order to provide systematic and comprehensive basic data support for subsequent evaluation work, the regional hydrogeological conditions include the distance between coal seams and aquifers, the lithological composition and thickness of aquitards, the strength of rock strata and the characteristics of initial fracture development, the distribution and scale of geological structures, the water-rich state of aquifers and water pressure, etc. To provide accurate and targeted detection data support for evaluating the water-blocking performance of the foundation, the geophysical exploration employs a combined approach of transient electromagnetic method and high-density resistivity method, focusing on detecting the distribution characteristics of adverse geological bodies such as concealed faults and densely fractured zones. The borehole exploration focuses on obtaining characteristics such as the distance between the coal seam and the aquifer, the lithological composition and thickness of the aquitard, the rock layer strength, and the development of initial fractures. Simultaneously, to ensure data reliability, the hydrogeological parameters in the detection results meet the following measurement accuracy requirements: aquifer water pressure error ≤ 0.05 MPa, rock layer thickness measurement error ≤ 5%, and rock layer strength test error ≤ 3%. Step 2: Based on geological conditions such as burial depth and aquifer water pressure, and mining conditions such as mining thickness, advance speed, and working face length, establish a composite floor plastic slip failure mechanical model under coal seam mining conditions. Clarify the floor plastic slip failure depth and state corresponding to different floor types, different failure states, and mining conditions. At the same time, conduct multi-dimensional field verification using borehole detection method, network parallel electrical resistivity tomography monitoring technology, and distributed optical fiber monitoring technology to determine the maximum development depth of the floor mining-induced failure zone. In order to accurately quantify the plastic slip failure depth of the base plate through a dedicated formula, with clear parameter definitions and a design that fits reality, and to support the accurate calculation of subsequent failure depth, the composite base plate plastic slip failure mechanics model includes three types of base plates: single-structure type, double-layer structure type, and soft-hard composite type. Among them, the double-layer structure type base plate includes three failure states, and the soft-hard composite type base plate includes five failure states. The model calculation parameters are calibrated through field geological survey data and laboratory rock sample test data. The plastic slip failure depth of the base plate is obtained according to formula (1). ; (1); In the formula, This represents the total thickness of the aquitard from the bottom of the coal seam to the top of the aquifer, in meters (m). , , , , The fitting parameters are for 5 different models; , , These are the internal friction angles of the bottom rock strata for single-striated, double-layered, and soft-hard composite types, respectively, in degrees. The crack width is expressed in mm. For crack roughness; This refers to the hydrostatic pressure of the water-containing material, expressed in MPa. Initial permeability of the rock formation, in meters. 2 ; Permeability of the rock strata after mining, in meters. 2 ; The rock mass compressive strength is expressed in MPa. This represents the maximum principal stress actually borne by the rock strata, expressed in MPa. , These represent the thicknesses of each rock layer in the double-layer structure and the soft-hard composite base plate, respectively, all in meters (m).

[0027] To accurately determine the maximum development depth of the mining-induced damage zone in the base plate, the following process was used for on-site verification combining borehole detection, network parallel electrical resistivity tomography (EPT) monitoring, and distributed fiber optic monitoring: S21: Drilling detection method: By arranging at least 3 monitoring boreholes, the borehole depth is not less than 1.2 times the predicted failure depth, and the borehole verticality error is ≤1°, the borehole inspection instrument is used to observe the crack penetration status at different depths in the base plate. S22: Network parallel electrical resistivity monitoring method: Determine the fracture development depth by the area of ​​abnormally high apparent resistivity. In the mining roadway, at least 4 measuring lines are arranged using a combination of borehole and roadway methods. Each measuring line is equipped with more than 16 electrodes. The electrode spacing is ≤4m, the apparent resistivity detection accuracy is ≤10Ω・m, and the data acquisition interval is ≤4h. S23: Distributed fiber optic monitoring technology: Fiber optic sensors are deployed vertically along the base plate with a sensor spacing of ≤5m, strain measurement accuracy of ≤10με, and data sampling frequency of ≥1Hz. Stress concentration depth is identified by strain abrupt change characteristics. S24: Determine the maximum development depth of the mining-induced damage zone of the base plate by combining data from three types of monitoring.

[0028] Step 3: Numerical simulation method is used to determine the distribution range of the mining-induced damage zone of the floor. Combined with the borehole water injection test method to obtain rock leakage data to determine the water conductivity of fractures, borehole inspection method to directly observe the development morphology of fractures, and transient electromagnetic method to detect the range of rock mass physical property changes, a multi-parameter collaborative verification is carried out in the whole space to accurately define the distribution range of the mining-induced damage zone of the floor. To accurately output data related to the distribution of mining-induced damage zones in the floor slab and provide reliable support for subsequent evaluation, the numerical simulation uses FLAC3D software to establish a fluid-structure interaction (FSI) model. The model dimensions are determined based on the working face strike, dip length, and floor slab detection depth, with a length × width × height of no less than 400m × 400m × 200m. The FSI model incorporates a stress-damage-seepage coupling model to characterize the evolution of rock mass permeability. A double-yield model is used to simulate the compaction characteristics of collapsed rock blocks in the goaf. The simulation step size is set to 5-10m. Displacement constraints are applied around the model and at the bottom, and an equivalent overlying stratum pressure (pressure value = average unit weight of the stratum × burial depth) is applied at the top. Actual water pressure is applied to the floor aquifer. The model calculation convergence criterion is an unbalanced force ratio ≤ 1 × 10⁻⁶. -5 The model outputs the longitudinal distribution range and lateral boundary of the mining-induced damage zone, and finally considers whether to retain it or not. Combining the evolution law of rock mass thermal stress, the thermal parameters can be corrected according to the actual rock strata type during the model calculation process to avoid the calculation error of the damage range caused by parameter deviation. To accurately define the distribution range of the base plate mining damage zone, the process of multi-parameter collaborative verification across the entire space is as follows: S31: Borehole water injection test method: Conduct water injection tests in sections in at least two monitoring boreholes in key areas such as the middle of the working face and near the coal pillars on both sides. The test section is 2-3m long. The water injection pressure is controlled at 0.5-1.0MPa using the constant pressure method. Record the leakage of rock mass per unit length. When the leakage is ≥5L / (min・m), it is judged as a section with developed water-conducting fractures. S32: Drilling inspection method: Using a drilling inspection instrument with a resolution of ≥0.1mm, continuous observation is carried out along the entire depth of the monitored borehole to record the fracture aperture, density and continuity status, and to divide the section into fracture-free section, micro-fracture section and densely fractured section. S33: Transient Electromagnetic Method: Employs a multi-turn small-loop transmitting device with a transmission frequency of 50-500Hz, a detection depth of not less than 1.2 times the predicted damage zone depth, and detection angles covering 0°, 45°, 90°, and 135°; data processing uses a regularized inversion algorithm with an inversion accuracy of ≤10%. The range of rock mass property changes is determined by the apparent resistivity difference (the difference between post-mining and pre-mining is ≥200Ω・m) (the larger the difference, the more severe the rock mass damage and the more developed the fractures). S34: By combining the overlapping areas of the numerical simulation results and the three types of verification data, the top boundary (coinciding with the bottom boundary of the mining-induced damage zone), bottom boundary, and lateral extension range of the bottom plate mining-induced damage zone are determined, so as to accurately define the distribution range of the mining-induced damage zone. The rules for defining the distribution range of the mining-induced damage zone in the floor are as follows: the top boundary is the bottom boundary of the mining-induced damage zone; the bottom boundary is the depth that meets any of the following conditions: the depth where the rock mass damage variable in the numerical simulation is ≤0.1, the depth where the leakage in the borehole water injection test is <1L / (min・m), the depth where there are no obvious cracks when the borehole is inspected, and the depth where the transient electromagnetic apparent resistivity difference is <50Ω・m; the lateral extension range is based on the working face boundary and extends to both sides for no less than 20m, or to the stress stability zone of the coal pillar.

[0029] Step 4: Combine numerical simulation to determine the mining stress path of the rock strata in the mining-induced damage zone, and convert the mining stress path into an equivalent laboratory loading path. Conduct seepage experiments on each rock stratum within the mining-induced damage zone under the mining stress path to clarify the permeability evolution characteristics of rock samples at different stress stages and establish mathematical models for the permeability evolution of rock samples at different stress stages. In order to obtain a complete four-stage stress path and clarify the key parameters of each stage, so as to provide accurate and comprehensive data support for the subsequent equivalent transformation of stress path, the mining stress path is obtained through the following numerical simulation method: monitoring points are set up in layers according to rock strata within the mining damage zone, with at least 3 monitoring points set up in each rock stratum and a spacing of 5m. The stress evolution data of the monitoring points during the working face advance process is extracted to obtain a complete four-stage stress path including the original rock stress stage, the advanced stress concentration stage, the mining pressure relief stage, and the stress environment adjustment stage, and to clarify the stress magnitude, stress change rate and duration of each stage. To accurately match the stress magnitude and equivalent amplified stress change rate between the field and the laboratory, and to ensure that the mechanical response of the laboratory loading path is consistent with that of the field mining stress path, thus providing a realistic loading basis for subsequent seepage experiments, the equivalent transformation follows the principles of stress evolution equivalence and time scale compression. The stress magnitude at each stage is kept consistent with the ratio coefficient (1:1) between the field stress and the strength of the laboratory rock sample, and the stress change rate is equivalently amplified according to the compression ratio (100-200:1) between the field stress evolution cycle and the laboratory test cycle, to ensure the consistency of the mechanical response between the laboratory loading path and the field mining stress path. To ensure accurate and reliable experimental results that closely reflect actual field conditions, the seepage experiment employed a ROCKTOP multi-field coupling test system. This system meets the following requirements: maximum axial load range ≥2000kN, confining pressure control accuracy ≤±0.1MPa, and seepage flow rate measurement accuracy ≤1×10⁻⁶. -8 m 3 / s, temperature control accuracy ≤±1℃, the temperature is kept constant (20±2℃) during the test to eliminate the influence of temperature on permeability; the rock samples are prepared by core sampling from each rock layer in the mining-induced damage zone, conforming to the standard of "Methods for Determination of Physical and Mechanical Properties of Coal and Rock", the sample size is Φ50mm×100mm, and 3 parallel samples are prepared for each group of rock layers to reduce dispersion; the confining pressure is set according to the confining pressure value of the corresponding rock layer in the field (8-16MPa), the permeation medium is deionized water, the permeation pressure is controlled at 0.5-5MPa, the axial stress, axial strain, confining pressure, permeation flow rate and other parameters are recorded in real time during the loading process, and the loading-holding-depressurization-reloading is completed in sequence according to the four stages of the mining-induced stress path; To provide solid data support for the subsequent establishment of the permeability evolution model, the permeability evolution characteristics are determined by calculating the permeability values ​​at each stress stage. The permeability calculation employs either the steady-state method or the transient pressure pulse method, where the permeability is ≤1×10⁻⁶. -17 m 2 The transient method was used to determine the trend of permeability with stress (increasing, decreasing, or stabilizing) and the critical stress threshold. The steady-state method required meeting the flow stability criterion (flow deviation ≤ 5% for three consecutive measurements). The transient pressure pulse method required recording the attenuation curves of upstream and downstream pressure over time, and calculating the permeability by fitting the attenuation curves. The permeability value at each stress stage was taken as the arithmetic mean of three parallel samples, with a relative standard deviation ≤ 15%. To accurately characterize the evolution of permeability at different stress stages and enhance the reliability and applicability of the model, the permeability evolution mathematical model uses stress parameters (principal stress, confining pressure, and deviatoric stress) as independent variables and permeability as the dependent variable. It employs exponential, power, or linear functions for fitting the model, with a goodness-of-fit R0. 2 ≥0.85, ensuring accurate characterization of the evolution of permeability at different stress stages.

[0030] In order to accurately fit the permeability evolution law under axial stress and confining pressure, a mathematical model of permeability evolution is established according to formula (2); (2); In the formula, k is the permeability, α and β are two different fitting parameters, σ1 is the axial stress in MPa, and σ3 is the confining pressure in MPa. Step 5: Calculate the development range of the pressure-bearing guide zone of the bottom plate by combining key parameters such as aquifer water pressure, width of original fractures inside the rock strata and roughness, and calibrate by combining on-site water level observation data to comprehensively determine the actual thickness of the pressure-bearing guide zone of the bottom plate within the mining area; To accurately calibrate the development range of the pressure-bearing riser zone on the bottom plate and ensure the reliability of the relevant parameter definitions, the on-site water level observation data calibration is achieved by arranging at least 3 calibration boreholes. The depth of the calibration boreholes is not less than 1.5 times the calculated riser height, and the bottom of the boreholes is located 5-10m above the top plate of the aquifer. The water level elevation inside the boreholes is measured using a water level measuring rope or a pressure sensor, with a measurement accuracy of ≤0.1m. Continuous observation is conducted for 3 hydrological cycles, and the actual riser height is calculated from the stable water level elevation.

[0031] To ensure the accuracy and reliability of the key parameters obtained, and to provide solid data support for subsequent calculations, these key parameters were acquired through a combination of on-site testing and laboratory analysis, as follows: S51: Aquifer water pressure is monitored by hydrological observation wells. At least two representative wells are selected and monitored continuously for more than 72 hours. The stable water pressure value is used as the basis for calculation. The monitoring accuracy is ≤0.05MPa. S52: The width of the original fractures inside the rock strata is observed by stereomicroscope after core drilling. No less than 3 sections are observed for each rock core, and no less than 10 fractures are measured in each section. The arithmetic mean is taken as the fracture width of the rock strata, and the measurement accuracy is ≤0.01mm. S53: The roughness of the original fractures inside the rock strata is obtained by scanning the fracture surface profile. A laser scanner is used to scan the fracture surface with a scanning resolution of ≤50μm. The roughness is characterized by the undulation coefficient of the profile curve, with the undulation coefficient ranging from 0.1 to 5.0μm.

[0032] In order to quickly and accurately solve the core parameters and provide a clear and reliable calculation basis for determining the actual thickness of the bearing-bearing guide zone, the development range of the bearing-bearing guide zone of the bottom plate is calculated according to formula (3). ; (3).

[0033] Step 6: Based on the actual thickness and permeability evolution characteristics of each rock layer, the weighted average method is used to calculate the equivalent water-blocking thickness of the base plate. The weighting coefficient is determined based on the ratio of the permeability of each rock layer to the initial permeability. Combined with the presence of tectonic structures and other hydrogeological conditions in the base plate, a three-level water-blocking performance evaluation standard (excellent, medium, and poor) is established to achieve a quantitative evaluation of the overall water-blocking performance of the base plate and a safety risk classification.

[0034] To accurately determine the equivalent water-blocking thickness of the waterproof layer, the calculation process for the equivalent water-blocking thickness of the base plate is as follows: S161: Determine basic parameters; execute steps 1 to 5 to determine the thickness of the mining-induced failure zone, mining-induced damage zone, original rock stress zone, and pressure-bearing conduction zone, as well as the evolution law of rock permeability in the mining-induced damage zone. S162: Quantification of mining-induced water resistance coefficient; introduction of mining-induced water resistance coefficient The water-blocking coefficient is used to quantify the change in the water-blocking characteristics of the rock strata under mining action. According to formula (4), the water-blocking coefficient is calculated by the ratio of the initial permeability of the rock strata to the permeability at a certain moment. : (4); In the formula, k0 is the initial permeability of the rock layer; k is the permeability of the rock layer.

[0035] S163: General Calculation of Equivalent Water-Blocking Thickness; The equivalent water-blocking thickness is used to characterize the total thickness of the rock layer with a mining-induced water-blocking coefficient of 1. For rock layers with reduced water-blocking capacity, the equivalent thickness is the product of the rock layer thickness and the mining-induced water-blocking coefficient. Based on this, a general calculation is performed according to formula (5) to obtain the equivalent water-blocking thickness of the aquitard. ; (5) In the formula, The number of rock strata between the bottom of the coal seam and the top of the aquifer; The water resistance coefficient for mining of each rock stratum; The thickness of each rock layer is given in meters (m). S164: Specific calculations based on the "lower four zones" theory; combining the lower four zones theory, all rock strata from the bottom of the coal seam to the top of the aquifer are divided into mining-induced failure zone, mining-induced damage zone, original rock stress zone, and pressure-bearing uplift zone, with corresponding thicknesses as follows: , , , Based on the water-blocking characteristics of each stratum, determine its equivalent water-blocking contribution coefficient: the water-blocking capacity of the mining-induced failure zone and the confined uplift zone is extremely low or completely lost, so set its mining-induced water-blocking coefficient. The mining-induced damage zone, although not destroyed by coal seam mining disturbance, has gradually developed internal fractures, resulting in varying degrees of decrease in water-blocking capacity. A mining-induced water-blocking coefficient is set based on the degree of decrease in water-blocking capacity. Located between 0 and 1; the original rock stress zone has not been disturbed by mining, and its water-blocking capacity has not changed. Assume its mining-induced water-blocking coefficient is... Pressure-bearing riser zone It does not have an effective water-blocking function, so let its mining-induced water-blocking coefficient be... It should be noted that if the advanced stress level does not cause the development of cracks in the bottom plate, it may lead to a decrease in permeability. In this case, the water-blocking coefficient of the mining in this area is greater than 1. If there are geological structures such as faults, collapse columns, and densely fractured zones in the bottom plate, the equivalent water-blocking thickness needs to be reduced by the range of the corresponding geological structures. Based on this, specific calculations are performed according to formula (6) to obtain the equivalent water-blocking thickness of the water-resistant layer. ; (6).

[0036] To achieve accurate quantitative evaluation of water-blocking performance and precise risk classification, the process of quantitative evaluation and safety risk classification of the overall water-blocking performance of the base plate is as follows: S261: Calculation of the bottom plate water inrush coefficient; Considering the characteristics of the water-resistant layer, the corrected bottom plate water inrush coefficient is calculated based on the equivalent water-resistant thickness according to formula (7). ; (7); In the formula, The aquifer pressure is expressed in MPa. S262: Water-blocking performance evaluation; Based on on-site engineering practice and water-blocking layer protection requirements, a three-level water-blocking performance evaluation standard is adopted, with the water inrush coefficient corresponding to the equivalent water-blocking thickness as the core indicator: water inrush coefficient ≤ 0.04MPa / m is "excellent", corresponding to low risk, and normal mining is possible; 0.04MPa / m < water inrush coefficient ≤ 0.06MPa / m is "medium", corresponding to medium risk, and enhanced monitoring measures are required; water inrush coefficient > 0.6MPa / m is "poor", corresponding to high risk, and treatment measures such as bottom grouting reinforcement and dewatering pressure reduction are required before mining; the thickness of the key water-blocking layer must meet the basic requirement of not less than 20m. If it is less than 20m, further enhanced treatment measures are required.

[0037] The following section provides a detailed explanation of the overall water-blocking performance evaluation method for the substrate based on permeability evolution, using specific engineering scenarios as examples. Figure 1 As shown, this embodiment uses coal seam floor water inrush risk assessment as an application scenario to verify the feasibility and practicality of the method of the present invention. I. Project Background The 4301 working face in a certain mine has a strike length of 720m and a dip length of 180m. It mines the No. 4 coal seam with a mining height of 6.25m and a coal seam burial depth of 600m. The aquitard on the floor of this working face has a typical hard-soft-hard composite structure, consisting of fine sandstone (6.75m), sandy mudstone (11.63m), and fine sandstone (14.24m) from top to bottom. The initial permeability of each rock layer is 3.6×10⁻⁶. -17 m 2 1.2×10 -16 m 2 3.8×10 -17 m 2 The bottom plate contains an Ordovician limestone aquifer with a stable water pressure of 6.0 MPa. The aquifer is 153.5 m vertically from the coal seam, and a 140 m depth was observed during the working face advance. 3 The phenomenon of water inrush on the bottom plate poses a significant risk of sudden water inrush. It is necessary to use the method of this invention to quantitatively evaluate the overall water-blocking performance of the bottom plate and provide technical guidance for water hazard prevention and control.

[0038] II. Implementation Steps Step S1: Acquisition of hydrogeological information and establishment of database; By collecting mine geological exploration reports and supplementing exploration data, it was determined that the main aquifers on the bottom of the coal seam in the 4301 working face include the K3 sandstone fracture aquifer, the Taiyuan Formation L2 limestone aquifer, and the Middle Ordovician Fengfeng Formation and Shangmajiagou Formation limestone aquifers. Among them, the K3 sandstone aquifer and the L2 limestone aquifer have relatively weak water-bearing capacity and do not have the conditions for large-scale water inrush. The Ordovician limestone aquifer has strong water-bearing capacity and high pressure value (stable water pressure of about 6 MPa), and is the main water hazard threat to the safe mining of the coal seam.

[0039] Geophysical exploration was conducted using a combination of transient electromagnetic method (80m depth) and high-density resistivity method, successfully identifying two concealed faults, F1 and F2, with F1 dip angles of 35° and F2 dip angles of 42°. The fracture width of both faults is 1-2m. Simultaneously, five core drilling holes were deployed to measure the thickness of each rock layer, uniaxial compressive strength (56MPa for fine sandstone and 32MPa for sandy mudstone), and initial fracture parameters (average fracture width 0.12mm, roughness fluctuation coefficient Rz=0.85). Furthermore, three hydrological observation wells were deployed to continuously monitor aquifer water pressure for 72 hours, determining the stable aquifer water pressure to be 6.0MPa. Based on all the collected data, a complete hydrogeological information database for the 4301 working face was established.

[0040] Step S2: Determine the range of the mining-induced failure zone in the bottom plate; Based on the geological conditions of the working face (coal seam depth 600m, water pressure of Ordovician limestone aquifer 6.0MPa) and mining parameters (mining height 6.25m, advance speed 3m / d), a mechanical model of plastic slip failure of soft and hard composite floor was established. The model calculation determined that the floor was in failure state three (satisfying H1<H0

[0041] ​The experiment was conducted using a multi-dimensional field verification approach, as follows: ① Three monitoring boreholes were set up with a depth of 30m (not less than 1.2 times the predicted failure depth), and the borehole verticality error was ≤0.8° (meeting the accuracy requirement of ≤1°). Observation was conducted using a borehole inspection instrument, and continuous through-cracks were observed within the depth range of 22.8-23.3m; ② Four parallel electrical measuring lines were set up in the return airway, with 16 electrodes on each line and an electrode spacing of 4m (meeting the requirement of ≤4). The maximum depth of the area with abnormally high apparent resistivity (resistivity value 700-1500Ω・m) was detected to be 23.5m; ③ Distributed fiber optic sensors were deployed in monitoring borehole #1 with a sensor spacing of 5m and a strain measurement accuracy of 5με (meeting the requirement of ≤10με). A strain abrupt change (strain value >50με) was observed at a depth of 23.2m, indicating that this depth is a stress concentration area. Based on the above three types of monitoring data, the maximum development depth of the mining-induced damage zone on the bottom plate of the working face was finally determined to be 23.2m.

[0042] Step S3: Determine the extent of the mining-induced damage zone; A fluid-structure interaction model was constructed using FLAC3D software. The model dimensions were 400m × 400m × 300m (meeting the requirement that the length × width × height should not be less than 400m × 400m × 200m). A stress-damage-seepage coupling model was embedded in the model to characterize the evolution of rock mass permeability. The compaction characteristics of the collapsed rock blocks in the goaf were simulated using a double yield model, with a simulation step size of 5m (meeting the requirement of 5-10m). The model boundary conditions were set as follows: displacement constraints were applied to the four sides and bottom; an equivalent overlying stratum pressure (pressure value = average unit weight of the stratum × coal seam burial depth) was applied to the top; and an actual stable water pressure of 6.0MPa was applied to the Ordovician limestone aquifer at the bottom. The model calculation convergence criterion was set to an unbalanced force ratio ≤ 1 × 10⁻⁶. -5 The simulation calculation shows that the range of the mining-induced damage zone is 23.2-48.5m, and the corresponding thickness of the mining-induced damage zone is 25.3m.

[0043] The verification work was carried out using a multi-method collaborative verification scheme, specifically as follows: ① Water injection tests were conducted in sections within two monitoring boreholes, with each test section being 2-3 meters long. The constant pressure method was used to control the water injection pressure at 0.8 MPa (meeting the requirements of 0.5-1.0 MPa). (MPa requirement), monitoring results show that the depth of the water-conducting fracture section with a leakage rate ≥5L / (min・m) is up to 23.2m, which is consistent with the bottom boundary of the mining-induced damage zone; ② Using a borehole inspection instrument with a resolution of 0.1mm (meeting the requirement of ≥0.1mm), continuous observation was carried out along the entire depth of the monitoring borehole. The observation results show that there are no obvious fractures below 48.2m, and the water-blocking performance is good; ③ The transient electromagnetic method was used for detection, with the transmission frequency set to 50-500Hz, the detection depth not less than 1.2 times the predicted damage zone depth, and the detection angle covering 0°, 45°, 90°, and 135°. The data processing adopted a regularized inversion algorithm (inversion accuracy ≤10%). The detection results show that the depth of the apparent resistivity difference between the post-mining and pre-mining periods is <50Ω・m is 48.5m. Based on the combined numerical simulation results and three types of verification data, the top boundary of the mining-induced damage zone was determined to be 23.2m (coinciding with the bottom boundary of the mining-induced damage zone), and the bottom boundary was determined to be 48.5m. The lateral extension range was based on the working face boundary and extended to both sides by no less than 20m, which is consistent with the numerical simulation results.

[0044] Step S4: Establishment of the mining-induced stress path and permeability evolution model; Using the FLAC3D numerical simulation model, monitoring points were set up at the top, middle, and bottom of each rock stratum (upper fine sandstone, sandy mudstone, and lower fine sandstone) within the mining-induced damage zone. At least three monitoring points were set up for each rock stratum, with a spacing of 5m between the monitoring points. Complete stress evolution data of each monitoring point during the working face advance were extracted to obtain the four-stage mining stress path: original rock stress stage (deflector stress 18MPa) → advanced stress concentration stage (deflector stress 65.8MPa) → mining stress relief stage (deflector stress 3.0MPa) → stress environment adjustment stage (deflector stress 17.8MPa). The stress magnitude, stress change rate, and duration of each stress stage were clarified.

[0045] Following the principle of equivalent transformation, the field-induced stress path was equivalently transformed into a laboratory loading path with a time-scale compression ratio of 100:1, ensuring the consistency of the mechanical response between the laboratory loading path and the field-induced stress path (stress magnitudes at each stage were kept consistent at a 1:1 ratio, and stress change rates were equivalently amplified at a 100:1 ratio). The ROCKTOP multi-field coupling test system was employed, which meets the following requirements: maximum axial load range ≥2000kN, confining pressure control accuracy ≤±0.1MPa, and seepage flow measurement accuracy ≤1×10⁻⁶. -8 m 3 The requirements for temperature control accuracy are ≤±1℃. During the test, the temperature is kept constant at 20±2℃ to eliminate the influence of temperature on the permeability.

[0046] Rock samples were prepared from cores taken from various rock strata within the mining-induced damage zone, conforming to the standard "Methods for Determining the Physical and Mechanical Properties of Coal and Rock". The sample size was uniformly Φ50mm×100mm, with three parallel samples prepared for each rock stratum to reduce test dispersion. The confining pressure was set at 12MPa (meeting the requirement of 8-16MPa) based on the corresponding confining pressure value of the rock strata in the field. Deionized water was used as the permeation medium, with the permeation pressure controlled at 2MPa (meeting the requirement of 0.5-5MPa). During the loading process, parameters such as axial stress, axial strain, confining pressure, and permeation flow rate were recorded in real time. The loading-holding-depressurization-reloading process was completed sequentially according to the four stages of the mining stress path. The actual permeability of each rock stratum after mining in the stress stabilization stage was measured as follows: Upper fine sandstone 9.72×10⁻⁶. -17 m 2 Sandy mudstone 4.56×10 -16 m 2 Lower fine sandstone 10.36×10 -17 m 2 .

[0047] Based on the above seepage experimental data, with stress parameters (principal stress, confining pressure, and deviatoric stress) as independent variables and permeability as the dependent variable, an exponential function was used for fitting to establish a mathematical model of permeability evolution for each rock layer. The goodness of fit of the models all satisfied R. 2 The requirement of ≥0.85 is as follows: Upper fine sandstone: k=3.6×10 -17 e -0.012σ 1 +0.025σ 3 (R) 2 =0.91) Sandy mudstone: k=1.2×10 -16 e -0.008σ1+0.031σ3 (R) 2 =0.88) Step S5: Determine the thickness of the pressure-bearing guide belt; The hydrostatic pressure-fracture propagation coupled model was used for calculation, with the following parameters substituted: confined water hydrostatic pressure p0 = 6.0 × 10⁻⁶. 6 Pa, crack width B = 0.12 × 10 -3 m, crack roughness f=0.85, the calculated development height of the pressure-bearing guide zone is 6.9m.

[0048] Three calibration boreholes were drilled to a depth of 12m (not less than 1.5 times the calculated lift height). The bottom of the boreholes was located 5-10m above the top of the Ordovician limestone aquifer. The water level elevation inside the boreholes was measured using a water level measuring rope with a measurement accuracy of ≤0.1m. Continuous observation was conducted for three hydrological cycles, and the actual lift height of 6.7m was calculated based on the stable water level elevation. The difference between the calculated and measured values ​​was 3%, less than 10%, indicating that the calibration was successful. The actual thickness of the pressure-bearing lift zone on the bottom plate of the working face was ultimately determined to be 6.8m.

[0049] Step S6: Calculation of equivalent water-blocking thickness and evaluation of water-blocking performance; Substituting the key parameters obtained above (thickness of mining-induced failure zone, thickness and permeability of mining-induced damage zone, thickness of original rock stress zone, thickness of pressure-bearing conduction zone, etc.) into the specific calculation formula (6) for the equivalent water-blocking thickness based on the lower four-zone theory, the equivalent water-blocking thickness of the aquitard is calculated. The calculation process is as follows: ; Based on the calculated equivalent water-blocking thickness, and combined with the formula (7) for calculating the floor water inrush coefficient, the corrected floor water inrush coefficient considering the characteristics of the floor rock formation and the damage effect of the mining-induced damage zone is calculated: ; Based on the established three-level water-blocking performance evaluation standard, the overall water-blocking performance of the working face floor was evaluated: the calculated modified floor water inrush coefficient was 0.0547 MPa / m, which is within the range of 0.03-0.06 MPa / m, corresponding to a water-blocking performance level of "medium" and a water inrush risk of medium risk. Mining work can be carried out normally, but enhanced monitoring measures need to be taken to strengthen the real-time monitoring of floor rock deformation, fissure development, and aquifer water pressure to prevent floor water inrush accidents. This verifies the practicality and feasibility of the method of this invention.

[0050] This invention provides a method for evaluating the overall water-blocking performance of a substrate based on permeability evolution. Through systematic process design and technological innovation, it has the following advantages: 1. Solid foundational data and accurate and reliable definition of key parameters: Through geophysical exploration, borehole exploration, and the summarization of existing geological data, the hydrogeological conditions of the study area are systematically obtained, and a complete hydrogeological information database is established, providing a reliable data foundation for subsequent accurate evaluation work and avoiding evaluation bias caused by fragmented basic information. Based on this, a collaborative verification system integrating mechanical models, numerical simulations, and multi-site testing was constructed for the core parameters during the mining process of the floor slab, achieving precise parameter definition: First, a composite floor slab plastic slip failure mechanical model was established under coal seam mining conditions, combining geological and mining conditions, clarifying the depth and state of floor slab plastic slip failure under different working conditions. Simultaneously, multi-dimensional field verification was conducted using borehole detection, network parallel electrical resistivity tomography (EPM) monitoring, and distributed fiber optic monitoring to determine the maximum development depth of the floor slab mining-induced damage zone. Second, numerical simulation methods were used to clarify the distribution range of the floor slab mining-induced damage zone. Combined with borehole water injection tests, borehole inspection, and transient electromagnetic methods, multi-parameter collaborative verification was conducted throughout the space to further accurately define the distribution of the mining-induced damage zone. Third, the development range of the floor slab pressure-bearing and lifting zone was calculated based on key parameters, and its actual thickness was determined through calibration using field water level observation data. The application of these multi-technical cross-verification methods effectively avoided the blind spots and errors of single detection methods, significantly improving the accuracy of key parameter definition and providing reliable technical support for subsequent evaluation work.

[0051] 2. An innovative stress path equivalence method was developed, accurately capturing the evolution law of permeability. This invention breaks through the limitations of traditional static evaluation, focusing on the dynamic coupling relationship between mining stress and the permeability of the base strata. It innovatively adopts a technical approach combining in-situ stress path equivalence transformation with laboratory seepage experiments: First, the actual mining stress path of the mining-damaged strata is determined through numerical simulation. Then, this in-situ mining stress path is equivalently transformed into a laboratory loading path. Seepage experiments are then conducted on each stratum within the mining-damaged zone under the mining stress path, accurately capturing the permeability evolution characteristics of rock samples at different stress stages and establishing corresponding mathematical models of permeability evolution. This process achieves precise alignment between in-situ mining conditions and laboratory experiments, overcoming the shortcomings of traditional experiments that are detached from the actual mining stress environment and lack accurate characterization of permeability evolution laws. It can truly reflect the dynamic changes in the water-blocking performance of the base strata during mining, providing core theoretical and experimental support for subsequent quantitative evaluation.

[0052] 3. The calculation process for equivalent water-blocking thickness is scientific, and the evaluation criteria are adapted to complex working conditions. Based on the permeability evolution model, the innovative method adopts a weighted average method based on the ratio of initial permeability to post-mining permeability, combined with the actual thickness and permeability evolution characteristics of each rock layer, to calculate the equivalent water-blocking thickness of the base plate. This calculation method emphasizes the water-blocking contribution weight of low-permeability rock layers, effectively overcoming the problems of traditional weighted average methods neglecting the differences in water-blocking capacity of different rock layers and the disconnect between the calculated results and the actual water-blocking effect. This ensures that the equivalent water-blocking thickness can truly reflect the overall water-blocking level of the base plate, significantly improving the scientific nature of the calculation. Simultaneously, considering specific hydrogeological conditions such as the presence of structural features in the base plate, a three-level water-blocking performance evaluation standard is established, achieving precise alignment between quantitative evaluation of the base plate's water-blocking performance and safety risk classification, providing a clear basis for developing targeted prevention and control measures for areas with different risk levels.

[0053] 4. Breaking through the limitations of traditional evaluation methods, this invention demonstrates significant engineering applicability and promotional value. It overcomes the limitations of existing technologies that rely on qualitative evaluation based on static geological parameters. Instead, it focuses on the dynamic evolution of the floor's water-blocking performance during mining operations. Through quantitative analysis, precise verification, and scientific calculation throughout the entire process, it achieves dynamic quantitative evaluation of the overall water-blocking performance of the floor under mining conditions and precise classification of safety risks. The evaluation results are more closely aligned with actual mining conditions, providing a scientific and reliable technical basis for safe coal seam mining under the threat of high-pressure water, effectively ensuring safe and efficient mine operation. Meanwhile, the method is highly operable and widely applicable: it has designed suitable technical processes for different floor types such as single structure, double-layer structure, and soft-hard composite type, and quantified the operation parameters and criteria of each step. It does not require complex special equipment and is easy to promote and apply in the field. It is especially suitable for the prediction of floor water hazard risk in deep coal seam mining under the threat of high-pressure water. It is highly targeted for the quantitative evaluation of the water-blocking performance of composite floor. It has important engineering significance and application value for liberating stagnant coal resources under the threat of high-pressure water and reducing the risk of floor water hazard prevention and control, and has broad prospects for promotion.

[0054] 5. The core value is highlighted, contributing to improved quality and efficiency in water hazard prevention and control. This invention ultimately achieves precise quantitative evaluation of the water-blocking performance of composite foundations threatened by confined water. By refining key parameters, evolution patterns, and differentiating evaluation standards, it effectively addresses the pain points of existing technologies, such as difficulty in adapting to complex geological conditions of composite foundations and insufficient reliability of evaluation results. Its established three-level evaluation standard can clearly define the water hazard risk level of different areas, providing support for the development of targeted treatment plans such as "normal mining, enhanced monitoring, and grouting reinforcement" on-site. This avoids the blind application of water hazard prevention and control measures, improving the effectiveness of water hazard prevention and control while reducing costs, achieving a dual improvement in safe mining and economic benefits, further highlighting the engineering practical value and promotional significance of the method.

[0055] This method enables precise quantitative evaluation of the overall water-blocking performance of the coal seam floor, and is particularly suitable for quantitative evaluation of the water-blocking performance of composite coal seams. It effectively addresses the problem that existing technologies struggle to accurately characterize the evolution of water-blocking characteristics of aquifers during mining, especially in complex geological conditions with composite coal seams, under the threat of high-pressure water. This makes it difficult to provide a scientific basis for evaluating pressurized mining. This invention has significant engineering practical value and broad application value for ensuring safe and efficient coal mining under the threat of high-pressure water, promoting the release of stagnant coal resources, and improving mine safety production levels.

Claims

1. A method for evaluating the overall water-blocking performance of a substrate based on permeability evolution, characterized in that, Includes the following steps: Step 1: Through geophysical exploration, borehole exploration, and the summarization and organization of existing geological data, systematically obtain the hydrogeological conditions of the study area and establish a complete hydrogeological information database; Step 2: Based on geological and mining conditions, establish a composite floor plastic slip failure mechanical model under coal seam mining conditions to clarify the floor plastic slip failure depth and state corresponding to different floor types, different failure states, and mining conditions; at the same time, conduct multi-dimensional field verification by combining borehole detection method, network parallel electrical resistivity tomography monitoring technology, and distributed optical fiber monitoring technology to determine the maximum development depth of the floor mining failure zone; Step 3: Use numerical simulation to determine the distribution range of the mining-induced damage zone of the floor, and combine borehole water injection test, borehole inspection and transient electromagnetic method to conduct full-space multi-parameter collaborative verification to accurately define the distribution range of the mining-induced damage zone of the floor. Step 4: Combine numerical simulation to determine the mining stress path of the rock strata in the mining-induced damage zone, and convert the mining stress path into an equivalent laboratory loading path. Conduct seepage experiments on each rock stratum within the mining-induced damage zone under the mining stress path to clarify the permeability evolution characteristics of rock samples at different stress stages and establish mathematical models for the permeability evolution of rock samples at different stress stages. Step 5: Calculate the development range of the pressure-bearing guide zone of the bottom plate by combining key parameters, and calibrate it by combining on-site water level observation data to comprehensively determine the actual thickness of the pressure-bearing guide zone of the bottom plate within the mining area; Step 6: Based on the actual thickness and permeability evolution characteristics of each rock layer, the weighted average method is used to calculate the equivalent water-blocking thickness of the base plate; combined with the presence of tectonic structures and other hydrogeological conditions of the base plate, a three-level water-blocking performance evaluation standard is established to achieve quantitative evaluation of the overall water-blocking performance of the base plate and safety risk classification.

2. The method for evaluating the overall water-blocking performance of a substrate based on permeability evolution according to claim 1, characterized in that, In step 1, the regional hydrogeological conditions include the distance between the coal seam and the aquifer, the lithological composition and thickness of the aquitard, the strength of the rock strata and the characteristics of the initial fracture development, the distribution and scale of the geological structure, and the water-rich state and water pressure of the aquifer. The geophysical exploration employs a combined approach of transient electromagnetic method and high-density resistivity method, focusing on detecting the distribution characteristics of adverse geological bodies. The borehole exploration focuses on obtaining the distance between coal seams and aquifers, the lithological composition and thickness of aquitards, and the strength and initial fracture development characteristics of rock strata. Simultaneously, the hydrogeological parameters in the exploration results must meet the following measurement accuracy requirements: aquifer water pressure error ≤ 0.05 MPa, rock strata thickness measurement error ≤ 5%, and rock strata strength test error ≤ 3%. In step 3, the numerical simulation uses FLAC3D software to establish a fluid-structure interaction model. The model size is determined according to the working face strike, dip length, and bottom plate detection depth, with a length × width × height of not less than 400m × 400m × 200m. The fluid-structure interaction model is embedded with a stress-damage-seepage interaction model to characterize the evolution of rock mass permeability characteristics. The collapsed rock blocks in the goaf are simulated using a double yield model to simulate compaction characteristics. The simulation step is set to 5-10m. Displacement constraints are applied around the model and at the bottom, equivalent overlying rock pressure is applied at the top, and actual water pressure is applied to the bottom aquifer. The convergence criterion for the model calculation is an unbalanced force ratio ≤ 1 × 10⁻⁶. -5 Output the longitudinal distribution range and lateral boundary of the mining-induced damage zone; In step 5, the calibration of the on-site water level observation data is achieved by arranging at least 3 calibration boreholes. The depth of the calibration boreholes is not less than 1.5 times the calculated lift height. The bottom of the boreholes is located 5-10m above the top plate of the aquifer. The water level elevation inside the boreholes is measured using a water level measuring rope or a pressure sensor with a measurement accuracy of ≤0.1m. The data is continuously observed for 3 hydrological cycles, and the actual lift height is calculated by taking the stable water level elevation.

3. The method for evaluating the overall water-blocking performance of a substrate based on permeability evolution according to claim 2, characterized in that, In step 2, the composite base plate plastic slip failure mechanics model includes three base plate types: single-structure type, double-layer structure type, and soft-hard composite type. Among them, the double-layer structure type base plate includes three failure states, and the soft-hard composite type base plate includes five failure states. The model calculation parameters are calibrated by field geological survey data and laboratory rock sample test data. The plastic slip failure depth of the base plate is obtained according to formula (1). ; （1）； In the formula, This represents the total thickness of the aquitard from the bottom of the coal seam to the top of the aquifer, in meters (m). , , , , The fitting parameters are for 5 different models; , , These are the internal friction angles of the bottom rock strata for single-striated, double-layered, and soft-hard composite types, respectively, in degrees. The crack width is expressed in mm. For crack roughness; This refers to the hydrostatic pressure of the water-containing material, expressed in MPa. Initial permeability of the rock formation, in meters. 2 ; Permeability of the rock strata after mining, in meters. 2 ; The rock mass compressive strength is expressed in MPa. This represents the maximum principal stress actually borne by the rock strata, expressed in MPa. , These represent the thicknesses of each rock layer in the double-layer structure and the soft-hard composite base plate, respectively, all in meters (m).

4. The method for evaluating the overall water-blocking performance of a substrate based on permeability evolution according to claim 3, characterized in that, In step 2, the process of conducting on-site verification of dimensions using borehole detection, network parallel electrical resistivity tomography (OTT) technology, and distributed optical fiber monitoring technology is as follows: S21: Drilling detection method: By arranging at least 3 monitoring boreholes, the borehole depth is not less than 1.2 times the predicted failure depth, and the borehole verticality error is ≤1°, the borehole inspection instrument is used to observe the crack penetration status at different depths in the base plate. S22: Network parallel electrical resistivity monitoring method: Determine the fracture development depth by the area of ​​abnormally high apparent resistivity. In the mining roadway, at least 4 measuring lines are arranged using a combination of borehole and roadway methods. Each measuring line is equipped with more than 16 electrodes. The electrode spacing is ≤4m, the apparent resistivity detection accuracy is ≤10Ω・m, and the data acquisition interval is ≤4h. S23: Distributed fiber optic monitoring technology: Fiber optic sensors are deployed vertically along the base plate with a sensor spacing of ≤5m, strain measurement accuracy of ≤10με, and data sampling frequency of ≥1Hz. Stress concentration depth is identified by strain abrupt change characteristics. S24: Determine the maximum development depth of the mining-induced damage zone of the base plate by combining data from three types of monitoring.

5. The method for evaluating the overall water-blocking performance of a substrate based on permeability evolution according to claim 4, characterized in that, In step 3, the process of full-space multi-parameter collaborative verification is as follows: S31: Borehole water injection test method: Conduct water injection tests in sections in at least two monitoring boreholes in the key area. The test section is 2-3m long. The constant pressure method is used to control the water injection pressure at 0.5-1.0MPa. Record the leakage of rock mass per unit length. When the leakage is ≥5L / (min・m), it is determined to be a section with developed water-conducting fractures. S32: Drilling inspection method: Using a drilling inspection instrument with a resolution of ≥0.1mm, continuous observation is carried out along the entire depth of the monitored borehole to record the fracture aperture, density and continuity status, and to divide the section into fracture-free section, micro-fracture section and densely fractured section. S33: Transient electromagnetic method: Employs a multi-turn small loop transmitter with a transmission frequency of 50-500Hz, a detection depth of not less than 1.2 times the predicted damage zone depth, and detection angles covering 0°, 45°, 90°, and 135°; data processing uses a regularized inversion algorithm with an inversion accuracy of ≤10%, and the range of rock mass property changes is determined by the apparent resistivity difference. S34: By combining the overlapping areas of the numerical simulation results and the three types of verification data, the top boundary, bottom boundary and lateral extension range of the mining-induced damage zone of the base plate are determined, so as to accurately define the distribution range of the mining-induced damage zone. The rules for defining the distribution range of the mining-induced damage zone in the floor are as follows: the top boundary is the bottom boundary of the mining-induced damage zone; the bottom boundary is the depth that meets any of the following conditions: the depth where the rock mass damage variable in the numerical simulation is ≤0.1, the depth where the leakage in the borehole water injection test is <1L / (min・m), the depth where there are no obvious cracks when the borehole is inspected, and the depth where the transient electromagnetic apparent resistivity difference is <50Ω・m; the lateral extension range is based on the working face boundary and extends to both sides for no less than 20m, or to the stress stability zone of the coal pillar.

6. The method for evaluating the overall water-blocking performance of a substrate based on permeability evolution according to claim 5, characterized in that, In step 4, the mining stress path is obtained through the following numerical simulation: monitoring points are set up in layers of rock strata within the mining damage zone, with at least 3 monitoring points set up in each rock stratum and a spacing of 5m. The stress evolution data of the monitoring points during the working face advance process is extracted to obtain a complete stress path of four stages, including the original rock stress stage, the advanced stress concentration stage, the mining pressure relief stage, and the stress environment adjustment stage, and the stress magnitude, stress change rate and duration of each stage are determined. The equivalent transformation follows the principles of stress evolution equivalence and time scale compression. The stress magnitude at each stage is kept consistent with the ratio coefficient between the field stress and the strength of the laboratory rock sample. The stress change rate is equivalently amplified according to the compression ratio between the field stress evolution cycle and the laboratory test cycle to ensure the consistency of mechanical response between the laboratory loading path and the field mining stress path. The seepage experiment employed a ROCKTOP multi-field coupling test system, which meets the following requirements: maximum axial load range ≥2000kN, confining pressure control accuracy ≤±0.1MPa, and seepage flow rate measurement accuracy ≤1×10⁻⁶. -8 m 3 / s, temperature control accuracy ≤±1℃, the temperature is kept constant during the test to eliminate the influence of temperature on permeability; the rock samples are prepared by core sampling from each rock layer in the mining-induced damage zone, the sample size is Φ50mm×100mm, and 3 parallel samples are prepared for each group of rock layers to reduce dispersion; the confining pressure of the test is set according to the confining pressure value of the corresponding rock layer in the field, the permeation medium is deionized water, the permeation pressure is controlled at 0.5-5MPa, the parameters are recorded in real time during the loading process, and the loading-holding-depressurization-reloading is completed in sequence according to the four stages of the mining-induced stress path; The permeability evolution characteristics are determined by calculating the permeability values ​​at each stress stage. The permeability calculation uses either the steady-state method or the transient pressure pulse method, where the permeability is ≤1×10⁻⁶. -17 m 2 The transient method is used; the steady-state method satisfies the flow stability criterion; the transient pressure pulse method requires recording the decay curves of upstream and downstream pressures over time, and calculating the permeability by fitting the decay curves. The permeability values ​​at each stress stage were taken as the arithmetic mean of three parallel samples, with a relative standard deviation ≤15%. The mathematical model for permeability evolution uses stress parameters as independent variables and permeability as the dependent variable, and employs exponential, power, or linear functions for fitting. The goodness-of-fit R-value of the model is [value missing]. 2 ≥0.85, ensuring accurate characterization of the evolution of permeability at different stress stages.

7. The method for evaluating the overall water-blocking performance of a substrate based on permeability evolution according to claim 6, characterized in that, In step 5, the key parameters are obtained through a combination of on-site testing and laboratory analysis, as follows: S51: Aquifer water pressure is monitored by hydrological observation wells. At least two representative wells are selected and monitored continuously for more than 72 hours. The stable water pressure value is used as the basis for calculation. The monitoring accuracy is ≤0.05MPa. S52: The width of the original fractures inside the rock strata is observed by stereomicroscope after core drilling. No less than 3 sections are observed for each rock core, and no less than 10 fractures are measured in each section. The arithmetic mean is taken as the fracture width of the rock strata, and the measurement accuracy is ≤0.01mm. S53: The roughness of the original fractures inside the rock strata is obtained by scanning the fracture surface profile. A laser scanner is used to scan the fracture surface with a scanning resolution of ≤50μm. The roughness is characterized by the undulation coefficient of the profile curve, with the undulation coefficient ranging from 0.1 to 5.0μm.

8. The method for evaluating the overall water-blocking performance of a substrate based on permeability evolution according to claim 7, characterized in that, In step 4, a mathematical model of permeability evolution is established according to formula (2); （2）； In the formula, k is the permeability, α and β are two different fitting parameters, σ1 is the axial stress in MPa, and σ3 is the confining pressure in MPa. In step 5, the development range of the bearing-bearing guide zone of the bottom plate is calculated according to formula (3). ; （3）。 9. The method for evaluating the overall water-blocking performance of a substrate based on permeability evolution according to claim 8, characterized in that, In step 6, the calculation process for the equivalent water-blocking thickness of the base plate is as follows: S161: Determine basic parameters; execute steps 1 to 5 to determine the thickness of the mining-induced failure zone, mining-induced damage zone, original rock stress zone, and pressure-bearing conduction zone, as well as the evolution law of rock permeability in the mining-induced damage zone. S162: Quantification of mining-induced water resistance coefficient; introduction of mining-induced water resistance coefficient The water-blocking coefficient is used to quantify the change in the water-blocking characteristics of the rock strata under mining action. According to formula (4), the water-blocking coefficient is calculated by the ratio of the initial permeability of the rock strata to the permeability at a certain moment. : （4）； In the formula, k0 is the initial permeability of the rock layer; k is the permeability of the rock layer; S163: General Calculation of Equivalent Water-Blocking Thickness; The equivalent water-blocking thickness is used to characterize the total thickness of the rock layer with a mining-induced water-blocking coefficient of 1. For rock layers with reduced water-blocking capacity, the equivalent thickness is the product of the rock layer thickness and the mining-induced water-blocking coefficient. Based on this, a general calculation is performed according to formula (5) to obtain the equivalent water-blocking thickness of the aquitard. ; （5） In the formula, The number of rock strata between the bottom of the coal seam and the top of the aquifer; The water resistance coefficient for mining of each rock stratum; The thickness of each rock layer is given in meters (m). S164: Specific calculations based on the lower four-zone theory; combining the lower four-zone theory, all rock strata from the bottom of the coal seam to the top of the aquifer are divided into mining-induced failure zone, mining-induced damage zone, original rock stress zone, and pressure-bearing uplift zone, with corresponding thicknesses as follows: , , , Based on the water-blocking characteristics of each stratum, determine its equivalent water-blocking contribution coefficient: the water-blocking capacity of the mining-induced failure zone and the confined uplift zone is extremely low or completely lost, so set its mining-induced water-blocking coefficient. ; The mining-induced damage zone, although not destroyed by coal seam mining disturbance, has gradually developed internal fractures, resulting in varying degrees of decrease in water-blocking capacity. A mining-induced water-blocking coefficient is set based on the degree of this decrease. Located between 0 and 1; the original rock stress zone has not been disturbed by mining, and its water-blocking capacity has not changed. Assume its mining-induced water-blocking coefficient is... Pressure-bearing riser zone It does not have an effective water-blocking function, so let its mining-induced water-blocking coefficient be... ; Based on this, the equivalent water-blocking thickness of the waterproof layer is obtained by performing specific calculations according to formula (6). ; （6）。 10. The method for evaluating the overall water-blocking performance of a substrate based on permeability evolution according to claim 9, characterized in that, In step 6, the quantitative evaluation and safety risk classification of the overall water-blocking performance of the base plate are as follows: S261: Calculation of the bottom plate water inrush coefficient; Considering the characteristics of the water-resistant layer, the corrected bottom plate water inrush coefficient is calculated based on the equivalent water-resistant thickness according to formula (7). ; （7）； In the formula, The aquifer pressure is expressed in MPa. S262: Water-blocking performance evaluation; Combining on-site engineering practice and water-blocking layer protection requirements, the three-level water-blocking performance evaluation standard is based on the water inrush coefficient corresponding to the equivalent water-blocking thickness as the core indicator: water inrush coefficient ≤0.04MPa / m is excellent, corresponding to low risk, and can be mined normally; A water inrush coefficient of 0.04 MPa / m to 0.06 MPa / m is considered medium risk, requiring enhanced monitoring measures. A water inrush coefficient of >0.6 MPa / m is considered poor risk, requiring remediation measures before further mining. The thickness of the critical water-blocking layer must meet the basic requirement of not less than 20m. If it is less than 20m, further enhanced remediation measures are required.

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