Dynamic determination method for roadway support based on coupling of mining damage and stress deflection

Abstract

The present application relates to the technical field of mining engineering and roadway support, and discloses a roadway support dynamic determination method based on mining damage and stress deflection coupling, which obtains initial static ground stress parameters of a roadway and an initial included angle of maximum horizontal principal stress; monitors actual cumulative microseismic energy during mining, and combines with indoor test microseismic energy parameters to calculate rock mass damage parameters representing the rock mass degradation degree; uses the damage parameters to correct initial mechanical parameters to construct dynamic rock mass mechanical parameters; inputs the parameters into a three-dimensional numerical model simulation to extract the maximum horizontal principal stress dynamic deflection angle and obtain the actual dynamic included angle; calculates the dynamic support resistance of the advanced section according to the actual dynamic included angle, compares it with the initial conventional support resistance without deflection to obtain a support deviation coefficient; and determines the corresponding support strength adjustment strategy based on the deviation coefficient. The present application integrates rock mass damage evolution and stress principal axis deflection into support resistance calculation, and provides objective and quantitative basis for dynamic adjustment of on-site support parameters.

Description

Technical Field

[0001] This invention relates to the field of mining engineering and roadway support technology, specifically a dynamic determination method for roadway support based on the coupling of mining-induced damage and stress deflection. Background Technology

[0002] In mining engineering, controlling the surrounding rock of roadways is a fundamental aspect of maintaining the stability of underground space, and rationally determining the roadway support resistance is a prerequisite for formulating support parameters and construction strategies. Currently, roadway support design is usually completed during the excavation stage, and its resistance calculation and parameter setting are mainly based on the initial static in-situ stress obtained from on-site tests, as well as the static mechanical parameters of the original rock in its intact state.

[0003] However, during the mining process, the concentrated stress caused by mining exerts a continuous dynamic disturbance on the surrounding rock in the advance section of the working face. This disturbance is accompanied by the release of microseismic energy, leading to the initiation and propagation of microcracks within the rock mass, resulting in varying degrees of damage and deterioration. This is objectively manifested as a weakening of mechanical parameters such as the rock mass's elastic modulus, cohesion, and internal friction angle. The differences in the degree of stiffness degradation of the rock mass in different parts of the mining-affected area further alter the original stress transmission path, causing a dynamic deflection of the direction of the maximum horizontal principal stress. Existing methods for determining roadway support fail to objectively quantify the evolution of rock mass damage caused by mining through on-site monitoring, and also fail to incorporate the changes in roadway stress boundary conditions caused by the deflection of the principal stress axis into the calculation model of support resistance.

[0004] This static assessment mechanism, based on an undisturbed state, leads to discrepancies between the calculated conventional support resistance and the actual stress state under the influence of mining in the advanced section of the working face. Because existing technology cannot provide dynamic support resistance data that takes into account the coupling effect of rock mass damage and stress deflection, and lacks quantitative assessment rules for the deviation between dynamic and initial static stress, on-site construction often relies solely on subjective experience to blindly adjust support strength when facing mining-affected roadways. The lack of objective thresholds for judging support resistance deviations means that on-site operations such as adding or removing anchor cables or adjusting the density of individual props lack data-driven quantitative guidance. Summary of the Invention

[0005] To address the shortcomings of existing technologies, this invention provides a dynamic determination method for roadway support based on the coupling of mining-induced damage and stress deflection. This method solves the problem that existing methods for determining roadway support resistance do not take into account rock mass damage and stress principal axis deflection caused by mining at the working face, resulting in a deviation between the design of dynamic support parameters for the advanced section and the actual stress state.

[0006] To achieve the above objectives, the present invention provides the following technical solution:

[0007] This invention provides a method for dynamic determination of roadway support based on the coupling of mining-induced damage and stress deflection, comprising the following steps: The initial static geostress parameters at different spatial locations during tunnel excavation, as well as the initial angle between the initial maximum horizontal principal stress and the tunnel axis, were obtained. During the mining of the working face, microseismic monitoring was carried out to obtain the actual cumulative microseismic energy. Combined with the microseismic energy parameters measured by the indoor mechanical failure test of intact rock samples, rock mass damage parameters characterizing the degree of rock mass deterioration under different advance distances in the advance section of the working face were calculated. The initial mechanical parameters of the complete rock mass are obtained, and the initial mechanical parameters are corrected using the rock mass damage parameters to construct dynamic rock mass mechanical parameters under different advance distances. The initial static geostress parameters and the dynamic rock mass mechanics parameters are input into a three-dimensional numerical model for mining simulation. The dynamic deflection angle of the maximum horizontal principal stress is extracted, and the corrected actual dynamic angle is calculated by combining the initial angle. Based on the actual dynamic angle and the geometric characteristics of the roadway cross section, the dynamic support resistance of the advance section is calculated, the initial conventional support resistance without considering the deflection of the stress principal axis is obtained, and the dynamic support resistance is compared with the initial conventional support resistance to obtain the support deviation coefficient. Based on the support deviation coefficient, the corresponding support strength adjustment strategy is determined.

[0008] Regarding the innovative principle of this invention: The stress concentration caused by mining operations at the working face induces the initiation and propagation of microcracks in the surrounding rock, accompanied by the release of microseismic energy. This invention quantifies the damage evolution process of the rock mass by establishing the ratio between accumulated microseismic energy and ultimate failure energy. The increase in internal damage fissures in the rock directly leads to the weakening of macroscopic mechanical parameters, manifested as a decrease in elastic modulus, cohesion, and internal friction angle, as well as an increase in Poisson's ratio.

[0009] Substituting spatially differentiated dynamic rock mechanics parameters into a three-dimensional Mohr-Coulomb damage constitutive model can objectively reflect the local stiffness degradation caused by rock mass deterioration. Because the degree of stiffness degradation varies across different parts of the mining-affected area, the stress transmission path changes accordingly, leading to a dynamic deflection of the direction of the maximum horizontal principal stress.

[0010] The deflection of the principal stress direction changes the original load distribution pattern. By constructing an angle influence factor that reflects the change of the angle, it is introduced into the calculation process of roadway support resistance. The calculated dynamic support resistance objectively characterizes the actual stress boundary conditions of the roadway under the influence of mining.

[0011] Preferably, the initial static geostress parameters include the maximum horizontal principal stress, the minimum horizontal principal stress, and the vertical stress; the specific method for obtaining the initial static geostress parameters at different spatial locations during tunnel excavation is to use the hydraulic fracturing method to conduct in-situ tests along the tunnel axis in segments.

[0012] In one specific embodiment, the microseismic energy parameters measured by the indoor mechanical failure test of the intact rock sample include the average maximum microseismic energy released by the intact rock sample in the indoor uniaxial compression failure test, and the maximum number of microseismic events generated. The process of determining the rock mass damage parameters includes: Based on the average maximum microseismic energy and the number of maximum microseismic events, a limit microseismic energy reference benchmark characterizing the complete failure state of the rock mass is constructed; the quantitative proportion of the actual cumulative microseismic energy relative to the limit microseismic energy reference benchmark is evaluated, and the quantitative proportion is used as the rock mass damage parameter.

[0013] Furthermore, the initial mechanical parameters include initial elastic modulus, initial Poisson's ratio, initial cohesion, and initial internal friction angle; the initial mechanical parameters are corrected using the rock mass damage parameters to construct dynamic rock mass mechanical parameters under different lead distances. The specific process includes: establishing a negative correlation attenuation mapping relationship between the initial elastic modulus, the initial cohesion, the initial internal friction angle and the rock mass damage parameters, and obtaining the dynamic elastic modulus, dynamic cohesion and dynamic internal friction angle at the corresponding lead distance; establishing a positive correlation growth mapping relationship between the initial Poisson's ratio and the rock mass damage parameters, and obtaining the increased Poisson's ratio at the corresponding lead distance.

[0014] Preferably, the three-dimensional numerical model adopts the Mohr-Coulomb damage constitutive model; the extraction of the dynamic deflection angle of the maximum horizontal principal stress, combined with the initial angle, to calculate the corrected actual dynamic angle is specifically as follows: Extract the dynamic deflection angle of the maximum horizontal principal stress at different advance distances in front of the working face in the three-dimensional numerical model relative to the initial direction, and algebraically superimpose the initial angle and the dynamic deflection angle to obtain the actual dynamic angle; wherein, with the roadway axis as the reference, the dynamic deflection angle in the clockwise deflection direction is taken as a positive value, and the dynamic deflection angle in the counterclockwise deflection direction, which is opposite to the clockwise preset deflection direction, is taken as a negative value.

[0015] In one specific embodiment, the process of calculating the dynamic support resistance of the advance section based on the actual dynamic included angle and the geometric characteristics of the roadway cross section includes: obtaining the stress concentration coefficient of the roadway cross section and the equivalent area of ​​the roadway cross section, as well as the magnitude of the initial maximum horizontal principal stress in the initial static geostress parameters. Based on the actual dynamic included angle, an included angle influence factor is constructed to reflect the weight of the principal stress direction on the stress state of the roadway; wherein, the smaller the actual dynamic included angle, the smaller the included angle influence factor; the dynamic support resistance is determined according to the combined effect of the roadway cross-section stress concentration coefficient, the initial maximum horizontal principal stress, the included angle influence factor, and the equivalent area of ​​the roadway cross-section.

[0016] Furthermore, the calculation method for the initial conventional support resistance is determined to be consistent with that for the dynamic support resistance; the only difference is that the actual dynamic angle is replaced with the initial angle.

[0017] Furthermore, the process of comparing the dynamic support resistance with the initial conventional support resistance to obtain the support deviation coefficient is as follows: calculate the percentage change of the dynamic support resistance relative to the initial conventional support resistance, and use the percentage change as the support deviation coefficient.

[0018] Preferably, the support strength adjustment strategy based on the support deviation coefficient specifically includes: setting a preset deviation threshold; determining to maintain the existing support strength when the absolute value of the support deviation coefficient is less than or equal to the deviation threshold; determining to implement a reinforcement support strategy when the support deviation coefficient is greater than the deviation threshold; and determining to implement a weakening support strategy when the support deviation coefficient is negative and the absolute value of the support deviation coefficient is greater than the deviation threshold.

[0019] In one specific embodiment, the reinforcement support strategy specifically includes installing additional anchor cables and simultaneously increasing the density of individual support pillars; the weakening support strategy specifically includes canceling the additional anchor cable installation and reducing the support density of individual support pillars; the deviation threshold is set to 20%.

[0020] This invention provides a method for dynamic assessment of roadway support based on the coupling of mining-induced damage and stress deflection. It has the following beneficial effects: 1. This invention obtains the actual cumulative microseismic energy during mining and quantifies its proportion with the ultimate microseismic energy reference obtained from laboratory tests to derive rock mass damage parameters. This allows for the correction of initial elastic modulus and other rock mass mechanical parameters to obtain dynamic rock mass mechanical parameters. This technical feature transforms the deterioration and evolution behavior of rock mass during mining into a physically quantifiable calculation result, overcoming the technical deficiency of traditional assessment methods that rely solely on initial static mechanical parameters, thus failing to characterize rock mass damage.

[0021] 2. This invention inputs dynamic rock mass mechanics parameters and initial static geostress parameters into a three-dimensional numerical model, extracts the dynamic deflection angle of the maximum horizontal principal stress to obtain the actual dynamic angle, and constructs an angle influence factor to calculate the dynamic support resistance of the advance section. This technical feature objectively takes into account the change in stress transmission path caused by the local stiffness degradation of the mining-induced rock mass, so that the calculated support resistance value directly corresponds to the actual stress state after the deflection of the principal stress axis in the advance influence zone of the working face.

[0022] 3. This invention obtains the support deviation coefficient by calculating the percentage change in dynamic support resistance relative to the initial conventional support resistance, and compares the absolute value of this coefficient with a preset deviation threshold to trigger specific judgment strategies for maintaining, reinforcing, or weakening the support. This technical feature establishes a complete set of closed-loop data-driven judgment rules, directly converting the resistance deviation results into executable on-site construction instructions such as installing additional anchor cables and adding or removing individual props, avoiding subjective blindness in the adjustment of roadway support parameters. Attached Figure Description

[0023] Figure 1 A flowchart of the method provided in the embodiments of the present invention; Figure 2 This is an evolution diagram of the deviation coefficient of the mining advance section support provided in an embodiment of the present invention. Detailed Implementation

[0024] The technical solutions of the embodiments of the present invention will be clearly and completely described below with reference to the accompanying drawings. Obviously, the described embodiments are only some embodiments of the present invention, and not all embodiments. Based on the embodiments of the present invention, all other embodiments obtained by those skilled in the art without creative effort are within the scope of protection of the present invention.

[0025] See attached document Figure 1 Coal mining operations will cause a redistribution of the stress field within the surrounding rock mass. Accompanying the concentration of mining-induced stress, micro-fractures developed within the surrounding rock will gradually expand into macroscopic rock mass damage, thereby causing a deflection of the principal axis of the stress field in the region. In conventional roadway support design processes, static geological data from the tunneling period are often used as constant input parameters, with little consideration given to the coupling effect of the deterioration of the rock mass's physical and mechanical properties during the mining period and the dynamic evolution of the principal stress direction.

[0026] Based on the aforementioned physical phenomena and engineering background, the method provided in this embodiment aims to establish a coupled analysis mechanism among static in-situ stress, dynamic rock mass damage, and principal stress axis deflection, thereby outputting an adaptive support strength adjustment strategy. Its overall implementation process includes the following steps: As a preferred approach, in this embodiment, step S100 is first executed to obtain the initial static geostress parameters at different spatial locations during tunnel excavation, as well as the initial angle between the initial maximum horizontal principal stress and the tunnel axis.

[0027] The selection of multi-dimensional static geostress as the basic input parameter is based on the physical principle that this parameter directly determines the initial mechanical boundary state of the rock mass before mining disturbance. This step establishes the static stress benchmark required for all subsequent dynamic support assessment calculations.

[0028] As the working face continues to advance, the system enters the dynamic parameter acquisition stage, namely, executing step S200, which involves carrying out microseismic monitoring during the working face mining period to obtain the actual accumulated microseismic energy. Combined with the microseismic energy parameters measured by the indoor mechanical failure test of intact rock samples, the rock mass damage parameters characterizing the degree of rock mass deterioration at different advance distances in the advance section of the working face are calculated.

[0029] The reason for choosing microseismic energy as the evaluation index is that it can objectively and directly reflect the physical kinetic energy released by internal rock fractures. Considering the matching between field monitoring data and spatial location, the actual accumulated microseismic energy collected needs to be aligned with the dynamic conditions at a specific lead distance in terms of time and spatial coordinates. By comparing the actual released energy in the field with the ultimate failure energy measured by indoor physical tests, the nonlinear internal damage can be transformed into a calculable engineering parameter.

[0030] The macroscopic bearing capacity of a rock mass is directly mechanically related to the degree of microscopic damage within it. Based on this principle, step S300 is executed to obtain the initial mechanical parameters of the complete rock mass. The initial mechanical parameters are then corrected using the rock mass damage parameters to construct dynamic rock mass mechanical parameters under different lead distances.

[0031] This step establishes a mathematical mapping relationship between the initial mechanical properties and the degree of rock mass deterioration. The dynamically decaying or variable physical property parameters obtained based on this mapping mechanism provide material property inputs that conform to the on-site deterioration trend for subsequent numerical evolution analysis.

[0032] After obtaining the mechanical parameters of spatiotemporal evolution, step S400 is further implemented, in which the initial static geostress parameters and the dynamic rock mass mechanical parameters are input into the three-dimensional numerical model for mining simulation, the dynamic deflection angle of the maximum horizontal principal stress is extracted, and the corrected actual dynamic angle is calculated in combination with the initial angle.

[0033] The formation of goaf in the working face forces the load of the overlying strata to shift to the solid coal on both sides, inevitably causing a deflection of the local principal stress. This step cross-couples the obtained static boundary conditions with the dynamic deterioration physical property parameters in a numerical simulation environment to reproduce the deterioration process of the stress environment, thereby outputting key angle variables that can truly reflect the stress direction characteristics of the advanced section.

[0034] Based on the data flow transmitted by each logic module in the early stage, the final execution step S500 is carried out. According to the actual dynamic angle and the geometric characteristics of the roadway cross section, the dynamic support resistance of the advanced section is calculated, the initial conventional support resistance when no stress deflection occurs is obtained, and the dynamic support resistance is compared with the initial conventional support resistance to obtain the support deviation coefficient. Based on the support deviation coefficient, the corresponding support strength adjustment strategy is determined.

[0035] To avoid the risk of biased judgments due to relying on a single absolute extreme value, this embodiment constructs a support deviation coefficient by calculating the relative change range. Based on this parameter and combined with a preset deviation threshold range, the degree of deterioration or mitigation caused by stress deflection to the roadway's stress environment is quantitatively assessed, and then a gradient safety adjustment command, including reinforcement, maintenance, or reduction, is output.

[0036] Based on the overall workflow described above, and as a preferred embodiment, the mechanism for obtaining the initial mechanical boundary conditions during tunnel excavation is explained in detail. This stage occurs before the start of the working face mining operation, when the surrounding rock mass is in a state of in-situ stress equilibrium. The parameters obtained at this time will serve as static baseline input data for subsequent dynamic mining evolution analysis. This acquisition stage is specifically detailed into the following sub-steps: Step S110: Plan the spatial layout of in-situ ground stress test points according to the tunnel design axis.

[0037] To ensure that the input parameters are representative of the engineering spatial scale and to avoid data distortion caused by local geological structures, the measuring points are arranged in segments at preset intervals along the direction of the tunnel to be measured. As a preferred implementation method, the preset interval is set to 200m.

[0038] In practice, fault zones and areas of abnormally fractured surrounding rock should be avoided. In-situ test boreholes should be drilled deep into intact rock strata in the roadway walls. The bottom of the test borehole should be set as the target spatial location for subsequent stress detection, ensuring that the detection area is within the original rock stress zone and avoiding interference from the unloading effect of roadway excavation.

[0039] Step S120: The initial static geostress parameters are obtained at the target spatial location using the hydraulic fracturing method.

[0040] In underground mining engineering, the maximum horizontal principal stress is usually the dominant mechanical factor inducing asymmetric deformation and shear failure in the surrounding rock of the working face. Therefore, it is necessary to accurately determine its magnitude and direction as the core input parameter for subsequent coupled models. In this embodiment, in-situ hydraulic fracturing tests are conducted using test boreholes. High-pressure fluid is injected into a designated test section in the borehole to induce fracturing in the borehole wall rock mass, promote fracturing, and maintain the fracturing.

[0041] During this process, key field-measured data such as initiation pressure, re-tension pressure, and instantaneous closure pressure are recorded. Based on the aforementioned measured data, the three-dimensional principal stress components of the current measuring point are extracted. That is, the calculated maximum horizontal principal stress, minimum horizontal principal stress, and vertical stress are jointly constructed to form the initial static geostress parameters characterizing the stress state of the original rock at the current spatial location. For the specific mechanical operations of borehole fracturing using hydraulic fracturing, the real-time acquisition method of pressure data, and the conversion process of solving the pressure curve into corresponding stress values, those skilled in the art can refer to the current rock mechanics testing industry standards for implementation.

[0042] Step S130: Determine the direction of the initial maximum horizontal principal stress at the target spatial location, and calculate the initial included angle accordingly.

[0043] After obtaining the aforementioned stress values, it is necessary to further clarify the spatial orientation characteristics of the principal stresses. Specifically, this can be achieved by using an impression tool or an in-hole ultrasonic camera to obtain the orientation of the hydraulic fracturing cracks on the borehole wall. Since the macroscopic crack surfaces generated by hydraulic fracturing are typically perpendicular to the direction of the minimum horizontal principal stress and parallel to the direction of the maximum horizontal principal stress, the initial maximum horizontal principal stress direction at the current target spatial location can be determined accordingly.

[0044] Subsequently, the axial azimuth data of the roadway to be measured, recorded by the mine surveying system, were retrieved. The geometric angle between the direction of the initial maximum horizontal principal stress and the roadway axis on the horizontal projection plane was calculated, and this geometric angle was defined as the initial angle between the initial maximum horizontal principal stress and the roadway axis.

[0045] Based on the implementation of steps S110 to S130 above, the scattered engineering exploration data distributed in the tunnel excavation section is transformed into the physical boundary conditions required for numerical simulation. The obtained initial static geostress parameters and the magnitude of the maximum horizontal principal stress within them, together with the measured initial angle, serve as the core static input variables for downstream mining damage assessment and three-dimensional numerical simulation, thereby substantially supporting the extrapolation and calculation of dynamic support resistance during the working face mining period.

[0046] After completing the initial static mechanical boundary condition survey, this embodiment further conducts a quantitative assessment of the deterioration process inside the rock mass during the working face mining. As the working face continues to advance, mining stress will induce microscopic fractures in the coal and rock mass ahead. This hidden deterioration process is difficult to obtain through conventional surface observation.

[0047] Based on the positive physical correlation between energy dissipation and rock mass damage, this embodiment selects microseismic energy as a characterization parameter. The quantitative characterization stage of mining-induced rock mass damage parameters is specifically detailed into the following sub-steps: Step S210: Conduct a complete indoor mechanical failure test on the rock sample, and measure and extract the microseismic energy parameters as an evaluation benchmark.

[0048] Intact rock cores were drilled from the unaffected area at the corresponding elevation of the working face. After processing into standard rock samples, indoor uniaxial compression failure tests were conducted. During this process, acoustic emission or microseismic acquisition equipment was used to record the energy release characteristics of the rock sample throughout the entire loading and fracture process. The average maximum microseismic energy released by the intact rock sample in the uniaxial compression failure test was extracted, and the number of maximum microseismic events generated when the sample reached the failure limit was counted. The product of the above two parameters represents the upper limit of the total energy that the rock mass of this specific lithology can release under a completely unstable failure state. Based on this, a reference benchmark for the ultimate microseismic energy characterizing the complete failure state of the rock mass was constructed.

[0049] Step S220: Conduct microseismic monitoring during the working face mining period to obtain the actual cumulative microseismic energy at a specific advance distance.

[0050] During mining operations, a microseismic sensor array is deployed along the advance section of the roadway to capture micro-fracture signals within the surrounding rock induced by mining in real time. Preferably, within a 50m advance distance of the transport roadway and return airway, a set of microseismic sensors is deployed every 10m along the roadway axis.

[0051] Given that the data recorded by the microseismic monitoring system is time-series data generated continuously over time, while the physical subject of subsequent support determination is the spatial cross-section, it is necessary to align the time-dimensional monitoring data with the spatial-dimensional mining conditions.

[0052] Specifically, by combining the daily advance footage of the working face recorded by the mine production system, the micro-fracture energy collected within a specific time window (preferably 12 hours) is accumulated and mapped to the corresponding advance distance section in spatial location, thereby obtaining the actual accumulated microseismic energy at that specific advance distance. For the specific spatial topology arrangement of the microseismic sensor array and the specific algorithm for calculating the source energy based on waveform signals, those skilled in the art can refer to the conventional operating procedures in the field of microseismic monitoring.

[0053] Step S230: Evaluate the quantitative ratio of actual accumulated microseismic energy to the ultimate microseismic energy reference benchmark, and calculate the rock mass damage parameters.

[0054] The actual energy data obtained at a specific advance distance on-site was quantitatively compared with the ultimate failure benchmark obtained from laboratory tests. This quantitative ratio was used as a rock mass damage parameter characterizing the degree of rock mass deterioration at the corresponding spatial location. The specific calculation logic corresponds to the following mathematical relationship: ; In the formula, For rock mass damage parameters at a specific advance distance; This represents the actual accumulated microseismic energy obtained after spatiotemporal alignment processing; The average maximum microseismic energy released from an intact rock sample; This represents the maximum number of microseismic events.

[0055] In the above calculation model, given that indoor destructive tests inevitably involve energy release and microseismic events, all parameters in the denominator are always greater than zero, ensuring the effectiveness and mathematical stability of the calculation division. Meanwhile, this rock mass damage parameter... The valid value range is limited to 0 to 1. 0 represents that the original rock is intact and undamaged, and 1 represents that the rock mass is completely broken.

[0056] In practical engineering applications, if the calculations obtained from specific measuring points are affected by complex geological noise interference at the site... If the value is greater than 1, then according to the rock mechanics limit state principle, the rock mass damage parameter of this spatial node is forced to be 1 to prevent calculation divergence when input into downstream numerical simulation.

[0057] After obtaining the rock mass damage parameters characterizing the degree of rock mass deterioration as described above, the construction mechanism of dynamic rock mass mechanical parameters is further explained in detail as a preferred embodiment. Mining activities cause nonlinear variations in the internal structure of the surrounding rock in the advanced section of the working face. Continuing to use constant material properties would lead to severe distortion in subsequent numerical simulations. Therefore, establishing a mathematical evolution mapping between static mechanical indices and dynamic damage states is crucial. The construction stage of this dynamic rock mass mechanical parameter attenuation model is specifically detailed as follows: Step S310: Obtain the initial mechanical parameters of the complete rock mass.

[0058] In this embodiment, intact rock mass samples undisturbed by mining were extracted and subjected to conventional triaxial compression tests in the laboratory. During this loading and failure process, the initial elastic modulus, initial Poisson's ratio, initial cohesion, and initial internal friction angle of the samples were measured and extracted. The physical basis for choosing triaxial compression tests to obtain the aforementioned parameters is that it can best reproduce the true triaxial geostress constraint state of deep rock masses, ensuring that the obtained initial mechanical parameters have reliable static mechanical representativeness. For the specific instrument loading procedures and basic parameter calculation procedures for conducting indoor triaxial compression tests, those skilled in the art can refer to the current rock mechanics testing specifications.

[0059] Step S320: The initial mechanical parameters are corrected using the rock mass damage parameters to construct dynamic rock mass mechanical parameters under different lead distances.

[0060] The development and connection of microfractures within the rock mass directly manifest in the macroscopic mechanical response as a weakening of the rock mass's bearing stiffness, a reduction in shear strength, and an enhancement of its lateral volumetric expansion deformation capacity. Based on this rock mechanics damage evolution mechanism, a negative correlation attenuation mapping relationship is established between the initial elastic modulus, initial cohesion, initial internal friction angle, and the aforementioned rock mass damage parameters. This yields the attenuated elastic modulus, attenuated cohesion, and attenuated internal friction angle at the corresponding lead distance.

[0061] A positive correlation growth mapping relationship between the initial Poisson's ratio and the rock mass damage parameters is established simultaneously to obtain the increased Poisson's ratio at the corresponding lead distance. The specific calculation process for constructing the above dynamic rock mass mechanical parameters corresponds to the following mathematical relationship: ; ; ; ; In the formula, , , , These are the initial elastic modulus, initial Poisson's ratio, initial cohesion, and initial internal friction angle obtained in step S310, respectively. , , , Using rock mass damage parameters The corrected dynamic elastic modulus, dynamic Poisson's ratio, dynamic cohesion, and dynamic internal friction angle; , , , These are the attenuation coefficients for the corresponding mechanical properties, and their values ​​are usually limited to between 0 and 1.

[0062] In actual numerical calculations, to ensure the physical rationality and computational robustness of the algorithm, physical boundary constraints must be applied to the calculation results of the above formulas. Specifically, the theoretical lower limits of the elastic modulus, cohesion, and internal friction angle under deep damage conditions must not be less than the residual strength parameters of the rock.

[0063] When the above negative correlation calculation results approach or fall below zero, they are forcibly truncated to the corresponding residual mechanical parameter values. Simultaneously, for the dynamically increasing Poisson's ratio with positive correlation, based on the theoretical constraints of continuum mechanics, its absolute upper limit threshold is set to 0.5.

[0064] The aforementioned boundary truncation treatment effectively avoids the occurrence of negative values ​​or non-convergence of mechanical parameters under extreme damage conditions, thereby ensuring the normal solution of the downstream numerical model.

[0065] Step S330: Based on the on-site in-situ test data, the attenuation coefficient in the mapping relationship is inverted and corrected.

[0066] In the above calculation model, due to the different lithologies exhibiting varying sensitivities during fracture evolution, it is necessary to clearly define the specific values ​​of each attenuation coefficient. To eliminate subjective biases arising from purely theoretical and empirical settings, test boreholes were drilled at corresponding positions in the advanced section of the working face, and in-situ mechanical penetration tests were conducted using a borehole penetrometer to obtain the actual in-situ bearing capacity indicators of the rock mass at the current advanced distance.

[0067] The on-site bearing capacity index is substituted into the above calculation model and fitted and compared with the theoretical values ​​derived from damage parameters. The specific values ​​of various attenuation coefficients corresponding to the geological characteristics of this mining area are determined through inversion calculation. The inversion correction mechanism, which combines on-site measured data, ensures the objective accuracy of the material property inputs to the downstream three-dimensional numerical model.

[0068] After completing the dynamic mapping of the mechanical parameters of rock mass deterioration, this section will describe in detail the cross-coupling mechanism of static geostress and dynamic damage parameters in a three-dimensional numerical environment, as a preferred embodiment. The mining operation at the working face is essentially a dynamic evolution process that leads to the redistribution of the stress field in the surrounding rock mass. To accurately capture the spatiotemporal variation characteristics of the principal stress direction during this process, it is necessary to fuse and extrapolate the multi-source parameters obtained in the previous stage within the same physical field. This stress deflection coupling simulation and dynamic angle correction stage is specifically detailed into the following sub-steps: Step S410: Construct a three-dimensional numerical model that includes the spatial features of the roadway and the longwall face, and apply an initial static mechanical boundary.

[0069] In this embodiment, a three-dimensional numerical grid model reflecting the spatial structure of the engineering entity is established based on the actual geological borehole data and mining design drawings of the mine area to be evaluated. In order to eliminate the boundary effect interference caused by artificially truncated boundaries on the calculation results, the spatial size of the model is set to a preset multiple of the characteristic size of the roadway and working face mining space.

[0070] Displacement constraints are applied to the lateral and horizontal boundaries of the model, and the initial static geostress parameters of the multi-dimensional spatial dimensions obtained in step S100 are used as the initial stress field and loaded onto the lateral boundary of the three-dimensional numerical model. The stress input value of the lateral boundary dynamically changes every 200m along the axial direction of the roadway, thereby restoring the original rock stress equilibrium state before the rock mass was disturbed by mining. For the specific mesh generation operation, mesh size gradual control, and calculation convergence control logic of constructing the three-dimensional numerical model mesh, those skilled in the art can use conventional large-scale commercial numerical simulation software.

[0071] Step S420: Set the constitutive relation of the model and input the dynamic rock mass mechanics parameters and the corresponding spatial grid.

[0072] In the three-dimensional numerical model, the Mohr-Coulomb damage constitutive model is selected for the surrounding rock material properties. The physical reason for choosing this specific constitutive model is that it can effectively describe the nonlinear strength decay characteristics of the rock mass under shear yielding state, and highly matches the dominant failure mode of the disturbed rock mass. Based on this, spatial coordinate matching logic is executed.

[0073] Specifically, taking the current working face advance position as the spatial origin, different advance distance intervals are divided in front of the working face. The dynamic rock mass mechanical parameters (including dynamic elastic modulus, dynamic Poisson's ratio, dynamic cohesion, and dynamic internal friction angle) calculated in step S300 at different advance distances are directly input into the constitutive model as the material mechanical properties of the corresponding spatial grid regions. Through this spatial partitioning and assignment method, the engineering status of the gradient deterioration of rock mass properties in the advance section ahead of the working face under the influence of mining is objectively reproduced within the numerical model.

[0074] Step S430: Simulate the evolution process of mining at the working face and extract the dynamic deflection angle of specific spatial nodes.

[0075] After loading the boundary conditions and dynamic mechanical properties, the excavation command of the working face element is executed within the three-dimensional numerical model to simulate the unloading effect caused by the mining advance. As the stress field redistribution calculation reaches a convergence state, the stress tensor of the solid coal and rock mesh element at a specific advance distance in front of the working face is extracted.

[0076] Based on the stress tensor, the spatial vector direction of the maximum horizontal principal stress under the current mining disturbance state is calculated. During this calculation, algorithm stability must be determined. If the extracted local element stress tensor exhibits biaxial isobaric characteristics within the horizontal plane (i.e., the difference between the two horizontal principal stresses approaches zero), the direction of the principal stress will present uncertainty in mathematical solution (i.e., repeated eigenvalues ​​lead to matrix singularity). To ensure the completeness and engineering rationality of the algorithm, a stress difference threshold is set. When the difference in horizontal principal stresses is less than the threshold, the dynamic deflection angle output by the previous adjacent spatial node is directly used to prevent jumps in the numerical model output.

[0077] By geometrically comparing the extracted direction of the maximum horizontal principal stress under disturbance with the direction of the initial principal stress before excavation simulation on the same horizontal projection plane, the dynamic deflection angle of the maximum horizontal principal stress at a specific advance distance relative to the initial direction is calculated.

[0078] Step S440: Algebraically superimpose the dynamic deflection angle with the initial included angle to calculate the corrected actual dynamic included angle.

[0079] The core physical objective of this step is to transform the simple relative deflection variable into an absolute force angle relative to the fixed engineering axis of the roadway, so that it can be directly referenced in the downstream support design.

[0080] The initial included angle, pre-determined in step S100, is extracted and algebraically superimposed with the dynamic deflection angle output from the aforementioned three-dimensional numerical model according to predetermined rules. This yields the actual dynamic included angle that truly reflects the current force direction characteristics of the advanced segment. The specific calculation logic corresponds to the following mathematical relationship: ; In the formula, For a specific lead distance The actual dynamic included angle at the location; The initial angle is the angle between the initial maximum horizontal principal stress and the roadway axis when it is not affected by mining. For a specific lead distance The dynamic deflection angle at that location.

[0081] In the aforementioned algebraic superposition calculation rules, a unified mathematical notation system for the pre-defined spatial reference system and directions is essential. This embodiment pre-determines that, with the tunnel axis as the reference, the dynamic deflection angle in the clockwise direction is taken as a positive value, while the dynamic deflection angle in the counter-clockwise direction (opposite to the pre-defined clockwise deflection direction) is taken as a negative value. This notation rule ensures that the calculated dynamic angle has a clear physical directional meaning, avoiding spatial ambiguity caused by simple numerical addition, and providing rigorous data support for accurately constructing the angle influence factor downstream.

[0082] After extracting the spatiotemporal evolution characteristics and dynamic mechanical parameters of the mining-induced stress field, this section will focus on explaining the specific logic of dynamic adjustment of support strength as a preferred method in this embodiment. The stress state of the surrounding rock in the advanced section of the working face is directly controlled by the deflection and concentration of the stress field. This stage transforms the actual in-situ stress parameters obtained in the early stage into a support control model that can be directly quantified and executed on the engineering site. The support deviation assessment and dynamic adjustment judgment criterion stage is specifically detailed into the following sub-steps: Step S510: Based on the corrected actual ground stress parameters and tunnel cross-sectional morphology, construct a dynamic support resistance calculation model.

[0083] The degree of effective deviatoric stress concentration caused by the maximum horizontal principal stress on the surrounding rock of the roadway is closely related to its direction of action relative to the roadway axis. To objectively quantify this spatial orientation effect, a support resistance assessment model is established that correlates ground stress, angle, and roadway spatial geometric characteristics. The actual dynamic angle obtained in step S400 is converted into an angle influence factor, and the dynamic support resistance at a specific advance distance is calculated in combination with cross-sectional parameters. The specific calculation logic corresponds to the following mathematical relationship: ; In the formula, For a specific lead distance Dynamic support resistance at the location; This is the stress concentration factor of the cross section. Its specific value is determined according to the cross-sectional shape of the roadway. For rectangular cross sections, the value ranges from 1.2 to 1.4, and for semi-circular arched cross sections, the value ranges from 1.0 to 1.2. The initial maximum horizontal principal stress obtained in step S100 is used. In this evaluation model, the deflection effect is mainly reflected by the included angle. The maximum horizontal principal stress is the initial static measurement value. This is the actual dynamic included angle. That is, the influence factor of the angle between the maximum horizontal principal stress and the roadway axis; The equivalent area of ​​the tunnel cross-section is expressed in m². 2 .

[0084] In the above physical model, the included angle influence factor characterizes the nonlinear transformation relationship between principal stress deflection and surrounding rock load. When the actual dynamic included angle... The smaller the value, the smaller the calculated result of this factor, which means that the shear failure load distributed along the cross-section of the roadway is lower, and thus the required support resistance and support strength are also smaller.

[0085] Step S520: Calculate the support deviation coefficient to quantify the impact of dynamic disturbance on the initial support state.

[0086] To objectively reflect changes in support requirements caused by mining disturbances and avoid relying solely on a single absolute load value, a relative change comparison mechanism is used to construct evaluation indicators. The initial conventional support resistance, without considering principal axis deflection, is set as... The calculation formula is as follows: ,in This is the initial included angle.

[0087] Extract the dynamic support resistance obtained in the above steps, calculate its percentage change relative to the initial conventional support resistance, and define it as the support deviation coefficient, corresponding to the following mathematical relationship: ; In the formula, This represents the support deviation coefficient. In deep mining engineering, the initial support resistance without considering disturbance is... Primarily used to resist the stress of the original rock's own weight, the effective value must be greater than zero. This engineering physics background ensures, at its underlying logic, that the denominator of the above division operation is never zero, eliminating the risk of mathematical singularities leading to calculation divergence.

[0088] Step S530: Implement a dynamic adjustment strategy for support strength based on the support deviation coefficient.

[0089] In a preferred embodiment, the preset deviation threshold is set to 20%. Based on the preset evaluation threshold and the calculated support deviation coefficient, a graded control logic is executed, and a corresponding on-site support execution plan is output. When determining the support deviation coefficient When the absolute value is less than 20%, it indicates that the stress deflection caused by mining is within the safety margin of the existing support system. The system outputs a control command to maintain the original state, meaning that the support strength of this advanced section does not need to change during the mining period.

[0090] When determining the support deviation coefficient When the stress exceeds 20%, it indicates that the mining-induced deviatoric stress has exceeded the initial support resistance. The system then forcibly outputs control commands to implement reinforced support. The specific engineering implementation includes: Based on the original support network, additional anchor cables were installed on the top and sides of the advanced section of the roadway, and individual support pillars were simultaneously and densely deployed to provide active high-strength shear stiffness.

[0091] When determining the support deviation coefficient When the value is less than -20%, it indicates that the principal stress axis deflection has caused this area to become a stress release zone or unloading zone. The system outputs control commands to implement a weakening support strategy. Its specific engineering implementation is as follows: The planned additional anchor cable installation in the advanced section was cancelled, and the density of individual support pillars in the roadway was appropriately reduced, thereby avoiding excessive waste of support materials while ensuring the stability of the surrounding rock.

[0092] To aid in understanding the technical solution of this invention and to verify the engineering effectiveness of the above-mentioned dynamic stress deflection calculation and support dynamic adjustment logic, a specific application example based on real coal mine geological data is provided below.

[0093] Specific application examples and effect verification: To further verify the engineering applicability of the method proposed in this invention, the scheme was applied to the longwall face of a deep, high-stress mine. The target roadway has a depth of 850m, a rectangular cross-section with a width of 4.5m and a height of 3.2m, and its equivalent cross-sectional area was calculated. It is 14.4m 2 Since the cross-sectional shape is rectangular, the stress concentration factor is selected. Complete coal and rock samples from the same stratum at the site were extracted and subjected to indoor triaxial compression tests to obtain complete rock mass mechanical parameters, specifically including the initial elastic modulus. initial Poisson ratio Initial cohesion Initial internal friction angle .

[0094] During the mining process, two typical working points (300m and 500m of the mining block) were selected for support resistance assessment and dynamic adjustment.

[0095] Working Condition 1: 300m from the mining site (high stress deflection zone of the working face) Initial geological and stress parameters were extracted at a depth of 300m from the working face. Based on in-situ test results, the initial maximum horizontal principal stress in this area was determined. Minimum horizontal principal stress Vertical stress The initial maximum horizontal principal stress and the angle between the roadway axis .

[0096] Based on the rock mass fracture characteristics calculated from the on-site microseismic monitoring data, the different lead distances in this area were quantified. Rock mass damage parameters Leading distance Inside, Take 0.4; Within the range, Take 0.3; Within the range, Take 0.2; Within the range, The value is set to 0.1. The mechanical parameter attenuation coefficient (matching) within the corresponding interval is obtained through on-site borehole mechanical penetration testing and inversion correction. and Attenuation model): Inside, to The values ​​are 0.6, 0.3, 0.5, and 0.4, respectively. Inside, to The values ​​are 0.58, 0.29, 0.47, and 0.38, respectively. Inside, to The values ​​are 0.56, 0.27, 0.45, and 0.36, respectively. Inside, to The values ​​are 0.53, 0.25, 0.44, and 0.34, respectively.

[0097] The aforementioned static geostress and dynamic rock parameters were substituted into a three-dimensional numerical model for multi-field coupled evolution simulation. The dynamic deflection angle of the principal stress axis at different advance distances during the mining process was extracted. : The inside is 40°. The inside is 30°. The inner angle is 15°. The inner angle is 5°. The deflection angle is then superimposed on the initial angle to obtain the corrected actual dynamic angle. And substitute it into the support resistance model to calculate the deviation coefficient. : exist Within the interval, the actual dynamic included angle Based on including The support model solution, support deviation coefficient ; exist Within the interval, the actual dynamic included angle Calculate the support deviation coefficient ; exist Within the interval, the actual dynamic included angle Calculate the support deviation coefficient ; exist Within the interval, the actual dynamic included angle Calculate the support deviation coefficient .

[0098] The system's judgment results show that the deviation coefficient within 50m ahead of this operating point is... All exceeded the 20% upper limit threshold. This indicates that the deflection of the mining-induced stress principal axis has caused this area to become a severe shear compression concentration zone. The system outputs a forced reinforcement command: reinforcement and support of varying degrees are required within a 50m advance range, and on-site operations include pre-installing long anchor cables and densifying individual support pillars.

[0099] Working Condition 2: 500m from the mining site (stress relief zone of the working face) Parameter extraction was performed at a depth of 500m from the working face to the mining level. The initial maximum horizontal principal stress in this area was determined. Minimum horizontal principal stress Vertical stress Initial angle .

[0100] Based on the same three-dimensional coupled model, the dynamic deflection angle of the principal stress axis affected by mining at this working point is extracted. It exhibits negative evolution: The interior is -40°. The interior temperature is -30°C. The interior is -25°. The inner angle is -15°. Perform angle correction and substitute it into the model calculation: exist Within the interval, the actual dynamic included angle Calculate the support deviation coefficient ; exist Within the interval, the actual dynamic included angle Calculate the support deviation coefficient ; exist Within the interval, the actual dynamic included angle Calculate the support deviation coefficient ; exist Within the interval, the actual dynamic included angle Calculate the support deviation coefficient .

[0101] The system's judgment results show that the deviation coefficient within 50m ahead of this operating point is... All values ​​are well below the -20% lower limit threshold. This indicates that the deflection of the principal stress axis actually reduces the adverse impact of the geostress direction on the support in this area. The system outputs cost reduction and optimization instructions: no additional anchor cables are needed for the advanced section 500m from the mining site, and the system also guides on-site construction to appropriately reduce the erection density of individual support pillars.

[0102] On-site experimental verification and effect comparison: See attached document Figure 2To further verify the engineering effectiveness of the above-mentioned dynamic support adjustment model, in this embodiment, a field comparison test was conducted simultaneously in the target mine using a traditional control section and the application section of this scheme.

[0103] In the conventional control section where this invention was not adopted, the on-site engineering design ignored the spatiotemporal variation characteristics of the principal stress axis deflection, and used a standard pre-support system with constant strength at both 300m and 500m of the working face. On-site monitoring data showed that when the working face advanced to 300m of the working face, due to the unpredicted stress concentration caused by the ultimate deflection, the roadway experienced severe spalling and localized support failure, leading to engineering instability. At 500m of the working face, although no significant mechanical damage occurred in the surrounding rock, the actual bearing pressure demand had decreased, and maintaining a constant support strength resulted in over-support.

[0104] In the test section, which contrasts with the engineering results, the model constructed in this invention objectively reflects the dynamic variations in the stress state. Combined with... Figure 2 The evolution trend shown indicates that, for condition one (300m from the mining level), the support deviation coefficient curve output by the system exceeded the +20% reinforcement threshold within a 50m lead range, and reached an extreme value of 212% within a 10m lead range. Based on this precisely quantified deviation evolution data, high-strength dynamic reinforcement operations with gradients were implemented on-site. Subsequent displacement monitoring showed that the deformation of the surrounding rock in this section was effectively controlled, avoiding support failure caused by misjudgment of the direction of ground stress, and effectively reducing the rate of roof collapse and spalling accidents.

[0105] against Figure 2 In the second working condition presented in the data, at the 500m mark of the mining area, the support deviation coefficient curve remained below the cost reduction and density reduction threshold of -20% throughout, and even dropped to -100% near the working face. Based on this, the system determined that the area was in a deflection and unloading state and output an execution command for cost reduction and density reduction to the site. After implementing this optimization strategy, the maximum deformation of the surrounding rock in the advanced section stabilized at 85mm, and compared with the traditional constant support method, the consumption of support materials in this section was substantially reduced by approximately 37.5%.

[0106] Based on the field verification results of the combined curve data, the coupled evaluation model established by this invention not only improves the safety of rock strata control in deep high-stress mines, but also achieves precise cost reduction and efficiency improvement of roadway support systems.

Claims

1. A method for dynamic assessment of roadway support based on the coupling of mining-induced damage and stress deflection, characterized in that, Includes the following steps: The initial static geostress parameters at different spatial locations during tunnel excavation, as well as the initial angle between the initial maximum horizontal principal stress and the tunnel axis, were obtained. During the mining of the working face, microseismic monitoring was carried out to obtain the actual cumulative microseismic energy. Combined with the microseismic energy parameters measured by the indoor mechanical failure test of intact rock samples, rock mass damage parameters characterizing the degree of rock mass deterioration under different advance distances in the advance section of the working face were calculated. The initial mechanical parameters of the complete rock mass are obtained, and the initial mechanical parameters are corrected using the rock mass damage parameters to construct dynamic rock mass mechanical parameters under different advance distances. The initial static geostress parameters and the dynamic rock mass mechanics parameters are input into a three-dimensional numerical model for mining simulation. The dynamic deflection angle of the maximum horizontal principal stress is extracted, and the corrected actual dynamic angle is calculated by combining the initial angle. Based on the actual dynamic angle and the geometric characteristics of the roadway cross section, the dynamic support resistance of the advance section is calculated, the initial conventional support resistance without considering the dynamic deflection angle is obtained, and the dynamic support resistance is compared with the initial conventional support resistance to obtain the support deviation coefficient. Based on the support deviation coefficient, the corresponding support strength adjustment strategy is determined.

2. The method for dynamic determination of roadway support based on the coupling of mining-induced damage and stress deflection as described in claim 1, characterized in that, The initial static geostress parameters include the maximum horizontal principal stress, the minimum horizontal principal stress, and the vertical stress. The specific method for obtaining the initial static geostress parameters at different spatial locations during tunnel excavation is as follows: The data were obtained by in-situ testing along the axial direction of the roadway using the hydraulic fracturing method.

3. The method for dynamic determination of roadway support based on the coupling of mining-induced damage and stress deflection as described in claim 1, characterized in that, The microseismic energy parameters measured by the indoor mechanical failure test of the intact rock sample include the average maximum microseismic energy released by the intact rock sample in the indoor uniaxial compression failure test, and the maximum number of microseismic events generated. The process of calculating the rock mass damage parameters includes: Based on the average maximum microseismic energy and the number of maximum microseismic events, a reference benchmark for the ultimate microseismic energy characterizing the complete failure state of the rock mass is constructed. The quantitative ratio of the actual cumulative microseismic energy to the ultimate microseismic energy reference benchmark is evaluated, and the quantitative ratio is used as the rock mass damage parameter.

4. The method for dynamic determination of roadway support based on the coupling of mining-induced damage and stress deflection as described in claim 1, characterized in that, The initial mechanical parameters include the initial elastic modulus, initial Poisson's ratio, initial cohesion, and initial internal friction angle. The initial mechanical parameters are corrected using the rock mass damage parameters to construct dynamic rock mass mechanical parameters under different lead distances. The specific process includes: Establish a negative correlation attenuation mapping relationship between the initial elastic modulus, the initial cohesion, the initial internal friction angle and the rock mass damage parameters, and obtain the dynamic elastic modulus, dynamic cohesion and dynamic internal friction angle at the corresponding lead distance; Establish a positive correlation growth mapping relationship between the initial Poisson's ratio and the rock mass damage parameter, and obtain the increased Poisson's ratio at the corresponding lead distance.

5. The method for dynamic determination of roadway support based on the coupling of mining-induced damage and stress deflection as described in claim 1, characterized in that, The three-dimensional numerical model adopts the Mohr-Coulomb damage constitutive model; The dynamic deflection angle of the maximum horizontal principal stress is extracted, and combined with the initial angle, the corrected actual dynamic angle is calculated, specifically as follows: Extract the dynamic deflection angle of the maximum horizontal principal stress at different advance distances in front of the working face in the three-dimensional numerical model relative to the initial direction, and algebraically superimpose the initial angle and the dynamic deflection angle to obtain the actual dynamic angle; Specifically, with the roadway axis as the reference, the dynamic deflection angle in the clockwise direction is taken as a positive value, and the dynamic deflection angle in the counterclockwise direction, which is opposite to the preset clockwise deflection direction, is taken as a negative value.

6. The method for dynamic determination of roadway support based on the coupling of mining-induced damage and stress deflection as described in claim 1, characterized in that, The process of calculating the dynamic support resistance of the advance section based on the actual dynamic angle and the geometric characteristics of the roadway cross-section includes: Obtain the stress concentration factor of the cross section and the equivalent area of ​​the roadway cross section, as well as the magnitude of the initial maximum horizontal principal stress in the initial static geostress parameters; Based on the actual dynamic included angle, an included angle influence factor is constructed that reflects the weight of the influence of the principal stress direction on the stress state of the roadway; The smaller the actual dynamic included angle, the smaller the included angle influence factor; The dynamic support resistance is determined based on the combined effect of the stress concentration factor of the cross section, the magnitude of the initial maximum horizontal principal stress, the included angle influence factor, and the equivalent area of ​​the roadway cross section.

7. The method for dynamic determination of roadway support based on the coupling of mining-induced damage and stress deflection as described in claim 6, characterized in that, The calculation method for obtaining the initial conventional support resistance is consistent with that for the dynamic support resistance; The only difference is that the actual dynamic included angle is replaced with the initial included angle.

8. The method for dynamic determination of roadway support based on the coupling of mining-induced damage and stress deflection as described in claim 1, characterized in that, The process of comparing the dynamic support resistance with the initial conventional support resistance to obtain the support deviation coefficient is as follows: Calculate the percentage change in dynamic support resistance relative to the initial conventional support resistance, and use the percentage change as the support deviation coefficient.

9. The method for dynamic determination of roadway support based on the coupling of mining-induced damage and stress deflection as described in claim 1, characterized in that, The support strength adjustment strategy based on the support deviation coefficient specifically includes: Set a preset deviation threshold; When the absolute value of the support deviation coefficient is less than or equal to the deviation threshold, it is determined that the existing support strength should be maintained. When the support deviation coefficient is greater than the deviation threshold, it is determined that a reinforcement support strategy will be implemented. When the support deviation coefficient is negative and the absolute value of the support deviation coefficient is greater than the deviation threshold, it is determined that a weakened support strategy will be implemented.

10. The method for dynamic determination of roadway support based on the coupling of mining-induced damage and stress deflection as described in claim 9, characterized in that, The implementation of the reinforcement support strategy specifically includes installing additional anchor cables and simultaneously increasing the density of individual support pillars. The implementation of the weakened support strategy specifically includes canceling the additional anchor cable installation and reducing the support density of individual support pillars; The deviation threshold is set to 20%.

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