A Dynamic Design Method and System for Lower Coal Seam Roof Cutting Parameters Based on the Fragmentation and Filling Effect

Abstract

This invention discloses a dynamic design method and system for roof cutting parameters in lower coal seams based on the effect of fractured expansion and filling. Specifically, it relates to the field of roadway surrounding rock control and rock cutting technology in mining engineering. It addresses the technical problems of difficult stability control and high maintenance costs in lower coal seam roadways during near-distance downward mining, due to the complex non-uniform stress field formed by the goaf and coal pillars of the upper coal seam. The method designs roof cutting parameters by determining the distribution characteristics of the non-uniform residual stress field and the bearing capacity characteristics of the immediate floor strata, and implements directional pre-splitting cuts. It monitors and evaluates the effectiveness of key causal links in the entire process chain from cut implementation to roof collapse and filling. The method dynamically corrects the roof cutting parameters based on the actual filling effect and process evaluation results. This achieves refined and dynamic optimization of roof cutting parameter design, effectively improving the stress environment of the roadway surrounding rock and enhancing the reliability of surrounding rock stability control by actively guiding and controlling roof fracture and fractured expansion filling along a predetermined path.

Description

Technical Field

[0001] This invention relates to the field of roadway surrounding rock control and rock cutting technology in mining engineering, and more specifically, to a dynamic design method and system for lower coal seam roof cutting parameters based on the effect of fracture and swelling filling. Background Technology

[0002] In the downward mining of closely spaced coal seams, the upper coal seam is usually mined first. The resulting goaf and reserved coal pillar sections create a complex and non-uniform residual stress field in the underlying strata. Against this backdrop, the layout and maintenance of roadways for mining the lower coal seam face significant challenges. Traditional solutions often involve re-excavating roadways within the lower coal seam, or using methods such as reserving wider coal pillars and implementing high-strength roadway backfilling to maintain the goaf-side roadways, aiming to avoid or resist stress transmission and influence from the overlying goaf and coal pillars.

[0003] In existing technologies, the surrounding rock of the lower roadway of the close-range coal seam is directly affected by the non-uniform residual stress field of the floor formed after the mining of the upper coal seam. The surrounding rock is in a high-stress or drastically changing environment for a long time, which leads to significant deformation and damage risks after the roadway is excavated. The stability control of the surrounding rock is extremely difficult. Existing roadway layouts and conventional support methods are difficult to effectively adapt to and cope with this complex stress environment induced by a specific mining sequence, resulting in high roadway maintenance costs and unstable results. Summary of the Invention

[0004] In order to overcome the above-mentioned defects of the prior art, the present invention provides a dynamic design method and system for the top cutting parameters of the lower coal seam based on the effect of fragmentation and filling to solve the problems mentioned in the background art.

[0005] To achieve the above objectives, the present invention provides the following technical solution:

[0006] The dynamic design method for roof cutting parameters of lower coal seams based on the fragmentation and filling effect includes the following steps:

[0007] S1: Determine the distribution characteristics of the non-uniform residual stress field formed in the floor strata of the upper coal seam after mining, and evaluate the structural bearing characteristics of the immediate floor strata of the lower coal seam.

[0008] S2: Based on the distribution characteristics of the non-uniform residual stress field and the structural bearing characteristics of the direct floor strata, design the roof cutting parameters of the lower coal seam roadway.

[0009] S3: Based on the designed roof cutting parameters, implement directional pre-splitting cuts in the roof of the lower coal seam roadway;

[0010] S4: Monitor and evaluate the effectiveness of key causal links in the chain reaction from the implementation of directional pre-splitting cuts to the collapse of the roof strata;

[0011] S5: After the lower coal seam working face is mined, based on the effectiveness of the key causal links in the chain process, the roof strata above the directional pre-splitting cut form the expected collapse pattern and collapse, forming broken and swollen gangue.

[0012] S6: Based on the actual effect of filling the goaf with crushed gangue and the evaluation results of the effectiveness of key causal links in the chain process, dynamically adjust the roof cutting parameters.

[0013] Furthermore, the distribution characteristics of the non-uniform residual stress field formed in the floor strata after the mining of the upper coal seam are determined, and the structural bearing characteristics of the immediate floor strata of the lower coal seam are evaluated, including:

[0014] In-situ stress data were obtained by drilling exploratory boreholes in the roof of the lower coal seam and measuring using the stress relief method.

[0015] Inversion calculations were performed based on coal seam geological conditions and in-situ geostress data to determine the distribution characteristics of the non-uniform residual stress field.

[0016] Mechanical parameters of the immediate floor strata were obtained by drilling core samples and conducting indoor rock mechanics tests.

[0017] Based on mechanical parameters and the results of borehole inspection of cored holes, the structural bearing characteristics of the direct bottom strata are evaluated.

[0018] Furthermore, based on the distribution characteristics of the non-uniform residual stress field and the structural bearing characteristics of the immediate floor strata, the roof cutting parameters of the lower coal seam roadway are designed, including:

[0019] Based on the distribution characteristics of the non-uniform residual stress field, stress concentration regions and stress relief regions are identified, and the distribution of stress relief regions guides the determination of the cutting position.

[0020] Based on the structural bearing characteristics of the direct bottom strata, the integrity and bearing capacity of the rock mass are determined, and the cutting height is determined based on the depth of the weak bearing capacity area.

[0021] Based on the combined cutting location and cutting height, determine the cutting parameters to be used for implementation.

[0022] Furthermore, when determining the cut-off height based on the structural bearing characteristics of the immediate floor strata, the specific steps include:

[0023] When the structural bearing characteristics of the direct bottom strata show the presence of weak interlayers or significant delamination, the design should be based on the principle of increasing the top cut height so that the collapsed rock mass can penetrate the weak layer.

[0024] When the structural bearing characteristics of the direct bottom strata show that the rock mass is intact and has strong bearing capacity, the design should be based on the principle of keeping the cut-off height within the range of the direct top.

[0025] Furthermore, based on the designed roof cutting parameters, directional pre-splitting cuts are implemented in the roof of the lower coal seam roadway, including:

[0026] Based on the location of the cut-off roof, mark the drilling location at the corresponding position on the roof of the lower coal seam roadway;

[0027] Determine the drilling depth and charge length based on the cut-off height;

[0028] Drill holes to the predetermined depth at the marked location and fill them with shaped charge blasting devices. Then, carry out directional pre-splitting blasting along the design direction to form directional pre-splitting cuts.

[0029] Furthermore, the effectiveness of monitoring and evaluating the key causal links in the chain reaction from the implementation of directional pre-splitting cuts to the collapse of the roof strata includes:

[0030] After directional pre-splitting, borehole television is used to detect the morphology of the cut surface and evaluate the quality of the cut surface formation.

[0031] Microseismic and stress monitoring devices were deployed on both sides of the cut surface to collect spatiotemporal data on stress redistribution and micro-fracture events under the influence of mining.

[0032] Based on spatiotemporal data, we analyze the changes in principal stress direction and the characteristics of microseismic event clustering to determine whether the fracture propagates directionally along the cut surface.

[0033] The effectiveness of key causal links is evaluated by combining the results of the assessment of the quality of the cut surface formation and the direction of fracture propagation.

[0034] Furthermore, when assessing the effectiveness of key causal links, the specific steps include:

[0035] Monitor the size and accumulation pattern of the collapsed roof behind the goaf to assess whether it forms effective support for the basic roof.

[0036] The results of the assessment of the collapse block size, accumulation pattern and the forming quality of the cut surface of the top plate are correlated to determine the integrity of the process chain.

[0037] Based on the judgment results of process chain integrity, the effectiveness of the causal links in the entire process from cutting to filling and support is comprehensively evaluated.

[0038] Furthermore, after the lower coal seam working face is mined, based on the effectiveness of the key causal links in the chain reaction process, the roof strata above the directional pre-splitting cuts form the expected collapse pattern and collapse, forming fragmented gangue, including:

[0039] Based on the assessment of the effectiveness of key causal links, the synergistic state of fracture propagation and roof rotation is determined.

[0040] After the coal seam working face is mined, under the guidance and coordination of directional pre-splitting cuts, the roof strata above the cuts fracture and collapse along the cut surface.

[0041] The collapsed rock mass broke apart within the goaf, and its volume expanded to form boulders that filled the goaf.

[0042] Furthermore, based on the actual effects of filling the goaf with crushed gangue and the evaluation results of the effectiveness of key causal links in the chain reaction process, the roof cutting parameters are dynamically adjusted, including:

[0043] Monitor the filling density and roof connection of crushed gangue after filling the goaf to obtain the actual filling effect;

[0044] Based on the evaluation results of actual filling effect and the effectiveness of key causal links, analyze the weak links and design deviations in the chain process;

[0045] Based on the analysis results, the cutting height or cutting position in the cutting parameters are corrected, and the corrected cutting parameters are used in the design and implementation of subsequent sections.

[0046] On the other hand, the present invention provides a dynamic design system for the roof cutting parameters of the lower coal seam based on the effect of fractured expansion and filling, comprising the following modules:

[0047] The stress assessment module is used to determine the distribution characteristics of the non-uniform residual stress field formed in the floor strata of the upper coal seam after mining, and to assess the structural bearing characteristics of the immediate floor strata of the lower coal seam.

[0048] The parameter design module is used to design the roof cutting parameters of the lower coal seam roadway based on the distribution characteristics of the non-uniform residual stress field and the structural bearing characteristics of the immediate floor strata.

[0049] The cutting module is used to implement directional pre-splitting cuts in the roof of the lower coal seam roadway according to the designed cutting parameters;

[0050] The process monitoring module is used to monitor and evaluate the effectiveness of key causal links in the chain reaction process from the implementation of directional pre-splitting cuts to the collapse of the roof strata;

[0051] The gangue generation module is used to, after the mining of the lower coal seam working face, based on the effectiveness of key causal links in the chain process, cause the roof strata above the directional pre-splitting cut to form the expected collapse pattern and collapse, forming crushed gangue.

[0052] The parameter adjustment module is used to dynamically adjust the roof cutting parameters based on the actual effect of the crushed gangue filling the goaf and the evaluation results of the effectiveness of key causal links in the chain process.

[0053] Compared with the prior art, the present invention has the following beneficial effects:

[0054] 1. Effective guidance and control of the roof failure and filling process under complex stress conditions in close proximity to coal seams have been achieved. By simultaneously analyzing the non-uniform residual stress field formed by the overlying goaf and the structural bearing characteristics of the underlying direct floor strata, the roof cutting parameter design has a dual scientific basis, significantly enhancing its pertinence. The continuous physical process of cut formation, fracture propagation, and collapse filling is managed as a monitorable and assessable chain process, and the effectiveness of its key causal links is used as the core evaluation index and dynamic adjustment basis. This transforms the entire technical solution from a one-time experience design into an adaptive system that can provide real-time feedback and continuous iteration. It can proactively adapt to the spatiotemporal variations of stress environment and rock mass structure in coal seam mining, changing the technical logic for dealing with such complex conditions.

[0055] 2. Significantly improved the stability control of the surrounding rock in the roadway. Through precise roof cutting design based on stress field and rock mass structure, combined with process monitoring to ensure that it collapses according to the predetermined pattern, the crushing and swelling filling effect of the collapsed rock mass can be efficiently utilized to form a stable support structure on the roadway sidewalls. This effectively resists the non-uniform stress transmission from the overlying goaf and coal pillars, keeping the roadway in a relatively decompression zone. This greatly enhances the reliability and predictability of technical measures. Real-time assessment of the effectiveness of process chain links allows potential problems to be identified and warned before the final filling effect appears, providing the possibility for timely intervention and reducing engineering risks. By dynamically optimizing the roof cutting parameters, repeated support or roadway repair work caused by improper design is reduced, improving mining efficiency while ensuring safety. Attached Figure Description

[0056] Figure 1 This is a flowchart of the dynamic design method for lower coal seam roof cutting parameters based on the fragmentation and filling effect of the present invention;

[0057] Figure 2 This is a schematic diagram of the structure of the dynamic design system for the top cutting parameters of the lower coal seam based on the effect of fragmentation and filling, according to the present invention. Detailed Implementation

[0058] 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 of ordinary skill in the art without creative effort are within the scope of protection of the present invention.

[0059] Example 1: Figure 1 The present invention provides a dynamic design method for the roof cutting parameters of the lower coal seam based on the effect of fractured expansion and filling, which includes the following steps:

[0060] S1: Determine the distribution characteristics of the non-uniform residual stress field formed in the floor strata of the upper coal seam after mining, and evaluate the structural bearing characteristics of the immediate floor strata of the lower coal seam.

[0061] S2: Based on the distribution characteristics of the non-uniform residual stress field and the structural bearing characteristics of the direct floor strata, design the roof cutting parameters of the lower coal seam roadway.

[0062] S3: Based on the designed roof cutting parameters, implement directional pre-splitting cuts in the roof of the lower coal seam roadway;

[0063] S4: Monitor and evaluate the effectiveness of key causal links in the chain reaction from the implementation of directional pre-splitting cuts to the collapse of the roof strata;

[0064] S5: After the lower coal seam working face is mined, based on the effectiveness of the key causal links in the chain process, the roof strata above the directional pre-splitting cut form the expected collapse pattern and collapse, forming broken and swollen gangue.

[0065] S6: Based on the actual effect of filling the goaf with crushed gangue and the evaluation results of the effectiveness of key causal links in the chain process, dynamically adjust the roof cutting parameters.

[0066] To determine the distribution characteristics of the non-uniform residual stress field formed in the floor strata of the upper coal seam after mining, and to assess the structural bearing characteristics of the immediate floor strata of the lower coal seam, the following steps were performed:

[0067] In the roof strata above the predetermined roadway location of the lower coal seam, exploratory boreholes are drilled perpendicular to the bedding direction. The layout of the exploratory boreholes needs to cover the projected area and influence boundary of the upper coal seam goaf and the remaining coal pillars. The boreholes can be laid out in a grid pattern, and the grid spacing is determined according to the complexity of the geological conditions. For example, a spacing of 20 to 50 meters can be used. Hollow-inclusion stress relief probes are installed at predetermined depths in each exploratory borehole. The installation depth of the probes is selected in the original rock stress zone outside the tunnel design outline, for example, in an area more than 5 meters deep outside the tunnel outline. The stress of the rock core outside the probe is relieved by advancing and rotating the drill bit, and the strain gauge inside the probe records the strain recovery data of the rock core. The probe with the recorded strain data is removed, and in the laboratory, according to the principle and calculation formula of the hollow-inclusion stress relief method, the measured strain value is converted into the magnitude and direction of the three-dimensional geostress at the measuring point, thereby obtaining the in-situ geostress data at that location. The obtained in-situ geostress data includes the magnitudes of the maximum principal stress, intermediate principal stress, and minimum principal stress, as well as the azimuth and dip angles of each principal stress in space.

[0068] Inversion calculations are performed based on coal seam geological conditions and obtained multi-point in-situ stress data to determine the distribution characteristics of the non-uniform residual stress field. Coal seam geological conditions specifically refer to the mining boundary of the upper coal seam, the size and spatial coordinates of the remaining coal pillars, and the thickness and lithological sequence of inter-layer strata. This information is derived from geological exploration reports and mine excavation engineering plans. The inversion calculation is based on elasticity theory, constructing a three-dimensional numerical model that includes the space of the upper coal seam goaf, the entity of the remaining coal pillar, the lower coal seam, and the inter-layer strata. Measured in-situ stress data are set as stress boundary conditions on the model boundary or as known stress constraints at specific coordinate points within the model. By iteratively adjusting the rock mechanics parameters or initial stress field assumptions in the model, the difference between the stress state calculated at the corresponding positions of the measured points within the numerical model after computational equilibrium and the measured in-situ stress data is minimized. This difference is typically quantified using the root mean square error (RMSE). The inversion calculation is considered convergent and iteration stops when the RMSE is less than 0.5 MPa. Once the iterative calculation meets the convergence condition, the stress state in the entire computational domain represented by the three-dimensional numerical model is the distribution characteristic of the determined non-uniform residual stress field. This distribution characteristic is output in the form of stress cloud map or data list, clearly showing the spatial location and stress value range of stress concentration area, stress relief area and stress gradient zone.

[0069] Mechanical parameters of the immediate floor strata were obtained by drilling and coring the strata and conducting laboratory rock mechanics tests. Coring boreholes were drilled at the designed floor locations in the lower coal seam roadway, penetrating the immediate floor strata and reaching the stable strata below. A dual-tube diamond drill was used to obtain complete core samples. The retrieved cores were numbered, sealed, and transported to the laboratory for processing into standard cylindrical specimens. Uniaxial compression tests were performed on the specimens to determine the uniaxial compressive strength and elastic modulus of the immediate floor strata; Brazilian splitting tests were conducted to determine the tensile strength of the immediate floor strata; and triaxial compression tests were performed under different confining pressures, such as 3 MPa, 6 MPa, and 9 MPa, to determine the peak strength of the rock samples. The cohesion and internal friction angle of the immediate floor strata were then obtained by fitting the results using the Mohr-Coulomb strength criterion. These test results together constitute the set of mechanical parameters of the immediate floor strata.

[0070] Based on the obtained mechanical parameters of the immediate floor strata and combined with the results of borehole inspection from the core drill, the structural bearing capacity of the immediate floor strata was assessed. Borehole inspection was conducted using a digital borehole camera. The camera probe was inserted into the core drill hole and then raised at a constant speed, continuously recording a sequence of images covering a 360-degree range of the borehole wall. By interpreting the image sequence, various structural planes existing in the immediate floor strata were identified, including bedding planes, fractures, and joints, and their depth, orientation, width, and presence of infill were recorded. The number of structural planes identified per unit borehole length, i.e., the linear density index, was used to quantify the degree of development of the structural planes. When assessing the structural bearing capacity, a comprehensive judgment is made based on mechanical parameters and borehole inspection results: If the mechanical parameters indicate high rock strength (e.g., uniaxial compressive strength greater than 30 MPa) and the linear density measured by borehole inspection is low (e.g., less than 3 lines per meter), then the direct floor rock strata are assessed as having an intact structure and strong bearing capacity. If the mechanical parameters indicate low strength, or although the strength index is acceptable, but the linear density measured by borehole inspection is high, indicating the presence of dense fractures or weak interlayers, then the rock strata are assessed as having a broken structure or containing weak interlayers, resulting in weak bearing capacity. If obvious and continuous separation fractures, i.e., delamination spaces, are observed between rock strata in the borehole inspection image, then regardless of the mechanical parameter values, significant delamination and severely degraded bearing capacity are assessed at this location. This assessment process ultimately outputs a qualitative judgment and quantitative description of the structural bearing capacity characteristics of different depth sections of the direct floor rock strata, and clearly indicates the depth location of the weak bearing capacity areas.

[0071] Based on the distribution characteristics of the non-uniform residual stress field determined in step S1 and the structural bearing characteristics of the immediate floor strata of the lower coal seam evaluated in step S1, the roof cutting parameters of the lower coal seam roadway are designed. The design process addresses the distribution characteristics of the non-uniform residual stress field; these characteristics exist in the form of stress cloud diagrams or data lists containing spatial coordinates and stress components, clearly defining the vertical and horizontal stress values ​​at different spatial locations. The method for identifying stress concentration areas and stress relief areas involves statistically analyzing the stress data of the entire analysis area to calculate its average stress value. A stress concentration factor threshold is set to define stress concentration areas. This threshold is based on engineering experience and statistical analysis of historical mine pressure manifestation data. For example, by analyzing the roadway deformation data and corresponding stress calculation values ​​after mining multiple similar working faces, it was found that when the vertical stress value exceeds 1.5 times the regional average vertical stress value, the probability of severe roadway deformation increases significantly. Therefore, 1.5 times is set as the stress concentration factor threshold, and the continuous spatial range that meets this condition is defined as the stress concentration area. A stress relief factor threshold is set to define stress relief areas. This threshold is based on the observation of the stress recovery law of the floor after the goaf stabilizes. For example, monitoring data shows that when the vertical stress value is less than 0.8 times the regional average vertical stress value, the rock strata movement tends to stabilize. Therefore, 0.8 times is set as the stress relief factor threshold, and the continuous spatial range that meets this condition is defined as the stress relief area. By applying these two thresholds, the boundary lines of stress concentration areas and stress relief areas are marked on the projection plane of the roadway direction and vertical direction. The principle for determining the top cutting position is to ensure that the designed top cutting line is located as close as possible to or parallel to the distribution range of the pressure relief zone, while striving to ensure that the end of the solid coal side of the top cutting line avoids the core part of the high stress concentration zone. In specific operation, the central axis or long axis of the pressure relief zone is used as the initial reference line for the top cutting position, and then fine-tuned in combination with the actual tunneling direction, so as to finally determine a top cutting position coordinate that is both in a relatively low stress environment and convenient for engineering implementation.

[0072] Based on the structural bearing characteristics of the immediate floor strata of the lower coal seam assessed in step S1, the integrity and bearing capacity of the rock mass are determined, and the cutting height is determined accordingly. The structural bearing characteristics of the immediate floor strata of the lower coal seam are given in the form of an assessment report, which includes information on the rock mass integrity classification, the location of weak interlayers, the location of significant delamination, and the depth of areas with weak bearing capacity. The basis for judging the integrity of the rock mass is the description of the degree of fracture development and the linear density of structural surfaces in the assessment report; a linear density threshold is set as the judgment criterion. This threshold is derived from the statistical summary of the results of extensive borehole inspections of stable and fractured rock strata. For example, it is found that the linear density of rock strata with good integrity is generally less than 3 fractures per meter, so 3 fractures per meter is set as the linear density threshold; when the linear density in the assessment report is lower than this linear density threshold, and the assessment report does not indicate the presence of weak interlayers or delamination, the rock mass is judged to be intact; when the linear density is higher than this linear density threshold, or the assessment report clearly indicates the presence of interlayers filled with weak material, the rock mass is judged to be fractured or contain weak interlayers. The assessment of bearing capacity integrates integrity information and mechanical parameters. A strength threshold is set as an auxiliary standard for judging bearing capacity. This threshold refers to the basic strength requirements for anchorable rock strata in the mine roadway support design specifications. For example, the uniaxial compressive strength of 30 MPa is usually used as an important reference point. For sections with intact rock mass and uniaxial compressive strength greater than this strength threshold, the bearing capacity is judged to be strong. For sections with fractures or interlayers, even if the strength data is acceptable, the bearing capacity is judged to have weak links. The core principle for determining the cut-off height is based on the depth of the weak bearing capacity area. Starting from the tunnel roof, analyze the rock strata layer by layer until the first stratum assessed as having strong bearing capacity and stable thickness is found. The bottom depth of this stratum can be used as a reference value for the basic cut-off height. If there is a clearly identified weak bearing capacity area above this depth, such as a weak interlayer with a thickness of 2 meters, the cut-off height must be increased based on the basic reference value to ensure that the designed collapse height of the rock mass can completely penetrate this weak area, allowing the collapsed rock fragments to accumulate and support the stable rock strata below the weak area. The increased value should be at least equal to the vertical thickness of the weak area, with a safety margin considered. This safety margin takes into account construction errors and the unevenness of the rock strata; for example, based on experience, an additional 0.5 to 1 meter may be added.

[0073] When the structural bearing characteristics of the immediate floor strata indicate the presence of weak interlayers or significant delamination, the design principle is to increase the cut-off height to allow the collapsed rock mass to penetrate the weak layer. Under this principle, the design process first determines the top and bottom burial depths of the weak interlayers or significant delaminations. The design cut-off height must be greater than the bottom burial depth of the weak interlayers or significant delaminations, ensuring that the rock column from the roadway roof to the top of the design cut-off height encompasses and extends beyond the entire weak structure. To determine a sufficient and economical extension value, the fragmentation characteristics of the collapsed rock mass need to be considered. The fragmentation coefficient of the collapsed rock mass is a key parameter, obtained through laboratory tests or engineering analogies. The laboratory test method involves crushing rock samples taken from the site in a press and measuring the volume ratio before and after crushing. The engineering analogy method involves referring to existing mining operations in the same mining area or mines with similar geological conditions. By back-analyzing the surface pressure observation data and roof collapse conditions, an empirical range of fragmentation coefficients is derived. For example, for the rock strata in this mining area, the fragmentation coefficient is estimated to be between 1.3 and 1.5. During the design, the cut-off height should meet the following conditions: the fragmentation volume obtained by multiplying the original volume of the rock mass below the designed cut-off height by the fragmentation coefficient should be able to effectively fill the goaf space enclosed by the cut-off height, mining height, and collapse angle, and provide roof support to the overlying rock strata. Through adjustment calculations, the final determined cut-off height should ensure that weak interlayers or significant delaminations fall completely within the designed collapse range, and that the fragmentation filling body can provide reliable support to the relatively stable overlying rock strata.

[0074] When the structural bearing characteristics of the immediate floor strata indicate that the rock mass is intact and has strong bearing capacity, the design should be based on the principle of maintaining the cut-off height within the immediate roof range. The immediate roof range refers to the total thickness of the rock strata located above the coal seam, capable of collapsing as mining proceeds after the goaf, and whose fragmented rock blocks can fill the goaf and provide initial support to the old roof. The design process begins by determining the lithological composition and total thickness of the immediate roof from the geological columnar section and assessment report. The design value for the cut-off height is capped at no more than the total thickness of the immediate roof. Further analysis is conducted to determine if there are significant strength differences within the immediate roof. If the strength of the lower and middle layers of the immediate roof is significantly lower than that of the upper layers, the cut-off height can be set near the strength variation interface to encourage the lower and middle layers to collapse fully. If the overall strength of the immediate roof is uniform, the cut-off height can be set as a proportion of the total thickness of the immediate roof. This proportion is set to balance full collapse with engineering economy. For example, numerical simulation analysis is used to analyze the filling and stress environment at different cut-off heights to select a proportion that meets the requirements and has a small value, such as 70% to 90% of the total thickness of the immediate roof. The core of this design principle is to guide the orderly collapse of the immediate roof through active cut-off, based on its natural collapse and breccia capacity.

[0075] By combining the cutting location and cutting height, the cutting parameters for on-site implementation are determined. The cutting location is a spatial line defined by a sequence of three-dimensional coordinate points, whose coordinates are derived from the results determined under the guidance of the pressure relief area distribution and have been converted to the mine measurement coordinate system. The cutting height is a numerical value derived from the results determined after analyzing the bearing characteristics of the immediate floor strata structure. Combining the spatial coordinates of the cutting location with the numerical value of the cutting height defines the specific cutting parameters for each construction site along the roadway direction. At a certain roadway mileage, the cutting parameters are defined as follows: the cutting surface is perpendicular to the roadway roof, its inner end is located at a point on the roof surface at a specific distance from the roadway coal wall, and its outer end extends to a specific vertical depth inside the roof, which is the cutting height, and the entire cutting surface extends along the design azimuth angle. These parameters are compiled into a construction parameter table, which includes data such as borehole starting coordinates, borehole azimuth angle, borehole inclination angle, and borehole depth, which can directly guide on-site operations, thus completing the design of the cutting parameters for the lower coal seam roadway roof.

[0076] Based on the cutting parameters determined in step S2, directional pre-splitting is carried out on the roof of the lower coal seam roadway. The drilling location is marked at the corresponding position on the roof of the lower coal seam roadway according to the cutting position. The cutting position is determined in step S2 as a spatial line defined by a sequence of three-dimensional coordinate points in the mine surveying coordinate system. Marking the drilling location involves projecting this spatial line onto the roof surface after the roadway has been actually excavated, using a total station or laser marker for on-site layout. The starting coordinates of the first designed borehole are read from the construction parameter table output in step S2. The corresponding planar position is found in the roadway using a surveying instrument, and a clear cross mark is made on the roof surface with paint. This mark is the center point of the borehole opening. Along the roadway direction, according to the borehole spacing specified in the construction parameter table, for example, every 1 meter or 0.8 meters, the center points of the openings of all subsequent boreholes are marked sequentially. Connecting all the marked points forms a construction trajectory line that coincides with the design cutting position line on the horizontal projection.

[0077] Based on the cut height determined in step S2, the borehole depth and charge length are determined. The cut height is a numerical value representing the distance vertically upward from the surface of the tunnel roof to the top boundary of the designed cut. The determination of the borehole depth must consider the borehole's inclination angle; when designed as a vertical borehole, the borehole depth equals the cut height; when designed as a borehole inclined at a certain angle to the roof normal, the length of the borehole along the borehole axis is calculated using trigonometric functions to ensure that the borehole's end reaches the position corresponding to the cut height in the vertical direction. For example, if the designed borehole has an angle of 10 degrees with the roof normal and a cut height of 5 meters, the borehole depth is approximately 5.08 meters. The charge length is determined based on the drilling depth, the reserved sealing length, and the effective shaped charge blasting requirements. The charge length is the portion of the drilling depth minus the sealing length. The sealing length is to ensure that the blasting energy acts inside the rock mass and prevent gas from escaping. Its setting is based on engineering safety regulations and rock properties. For example, the sealing length is set to be no less than 1.5 meters. Therefore, the charge length is equal to the drilling depth minus the sealing length. At the same time, the charge length must cover the entire target rock stratum section that needs to be pre-fractured, that is, from a certain depth below the top surface of the top plate to the top boundary of the cut.

[0078] Drill to the predetermined depth at the marked location and install the shaped charge blasting device. Perform directional pre-splitting blasting along the design direction to create a directional pre-splitting cut. Using a hydraulic anchor drilling rig or specialized drilling equipment, drill to the predetermined depth at the center point of the marked top plate borehole opening, according to the azimuth and inclination angles specified in the construction parameter table. During drilling, ensure the borehole wall is straight and clean rock dust from the borehole regularly. After drilling is complete, prepare the shaped charge blasting device; the shaped charge blasting device typically consists of a shaped charge tube, explosive, detonating cord, and detonator. The shaped charge tube is a specialized tube with shaped charge grooves in a specific direction on its wall. The direction of these grooves determines the concentrated release direction of the blasting energy; during installation, it is essential to ensure that the orientation of the shaped charge grooves is consistent with the direction of the designed cut surface. During loading, the detonating cord is inserted into the shaped charge tube, followed by the loading of the specified amount of emulsion explosive or other mining safety explosives. The assembled shaped charge blasting device is then slowly advanced into the borehole to the predetermined loading section. Finally, the borehole opening is tightly sealed with mud or a special sealing material, ensuring the sealing length meets the preset requirements. After all boreholes are loaded, a network connection is established according to the blasting design, connecting the detonators of each borehole to the main detonation line in series or parallel. After confirming that all personnel and equipment have evacuated to a safe area, the detonation command is issued. The blasting energy is released in a concentrated manner along the shaped charge groove, generating a strong tensile stress wave in the rock mass, preferentially forming and expanding cracks in the designed plane direction. Through one or multiple blasts in groups, a continuous, essentially straight pre-splitting crack, i.e., a directional pre-splitting cut, is formed in the roadway roof along the designed trajectory. This cut weakens the roof strata at a predetermined location and height. The parameters for directional pre-splitting blasting need to be adjusted; the explosive consumption per unit volume refers to the amount of explosive consumed per unit volume of rock. Its initial value can be referenced from blasting experience data of similar rock types. For example, for hard sandstone, it can be referenced to the range of 0.3 kg to 0.4 kg per cubic meter. Then, through small-scale test blasts, the cutting effect can be observed and fine-tuned by increasing or decreasing the value; the decoupling coefficient refers to the ratio of the borehole diameter to the explosive cartridge diameter. Its selection affects the blasting force characteristics. Usually, a range of values ​​is selected, such as 1.5 to 3, to form an effective splitting effect.

[0079] The effectiveness of monitoring and evaluating key causal links in the chain reaction from the implementation of directional pre-splitting cuts to the collapse of the roof strata is assessed. After the implementation of directional pre-splitting cuts, borehole television is used to detect the morphology of the cut surface and evaluate the quality of the cut surface formation. Specifically, a borehole television probe is lowered near an existing pre-splitting borehole or in a newly constructed inspection borehole parallel to and closely spaced with the pre-splitting borehole. The probe moves at a constant speed within the borehole, capturing 360-degree circumferential images of the borehole wall to obtain a continuous sequence of borehole wall images. By analyzing the image sequence, the trace characteristics of the cut surface exposed in the borehole are identified, including the continuity, width, and roughness of the traces. When evaluating the quality of the cut surface formation, crack width thresholds, continuity thresholds, and roughness evaluation criteria are set. The crack width threshold is set based on laboratory rock mechanics tests and numerical simulations. Tests show that when the average width of the open crack in the rock mass is greater than 3 mm, its effect on blocking and guiding stress waves is significantly enhanced. Therefore, a crack width threshold of 3 mm is set. During evaluation, the average width of the identified cut traces is compared with the crack width threshold of 3 mm. If the average width is greater than 3 mm, the cut opening is considered good. The continuity threshold is set based on the proportion of continuous length required to effectively cut off the stress transmission path. Through engineering analogy and statistics, a continuity threshold of 10% is set. During evaluation, the percentage of the cumulative length of discontinuous segments of the cut trace within the detection range to the total detection length is calculated and compared with the continuity threshold of 10%. If the percentage is less than 10%, the cut continuity is considered good. The roughness evaluation standard is a qualitative observation of whether the rock mass on both sides of the trace is flat and whether there are significant faults or fracture zones, serving as an auxiliary judgment criterion. Based on the above comparisons and evaluation results, the cut surface forming quality is assessed as excellent, good, medium, or poor.

[0080] Microseismic and stress monitoring devices are deployed in the rock mass on both sides of the cut face to collect spatiotemporal data on stress redistribution and microfracture events under the influence of mining. The microseismic monitoring devices are deployed by drilling to a certain depth in the tunnel roof and sidewalls, placing microseismic sensors deep within the boreholes and coupling them with a couplant. The sensor array covers a certain distance in front of and behind the cut face, for example, 30 meters in front of and behind the cut face. The sensor spacing is set according to the required monitoring accuracy, for example, 10 to 15 meters. The stress monitoring device uses borehole stress gauges, installed in the solid coal seam and roof strata on both sides of the cut face to monitor the dynamic changes in vertical and horizontal stress. The data acquisition system continuously records the vibration signals received by the microseismic sensors and the electrical signals output by the stress gauges. The microseismic signals undergo preprocessing, including filtering and noise reduction, picking up the arrival time of the first arrival wave, and calculating the source location and energy. The filtering frequency band in the preprocessing is determined based on the rock mass properties and environmental noise; for example, a bandpass filter is set to retain useful signals from 100 Hz to 1500 Hz. Source localization employs a three-dimensional localization algorithm based on arrival time difference. The input parameters are the coordinates of each sensor and the propagation speed of sound waves in the rock mass; the output is the spatial coordinates of the microseismic event. Stress data is preprocessed, including converting electrical signals into stress values ​​and eliminating the influence of temperature drift on sensor readings. The final spatiotemporal data includes the occurrence time of the microseismic event, its three-dimensional spatial coordinates, released energy, and stress value variation curves at different locations over time.

[0081] Based on spatiotemporal data obtained from microseismic and stress monitoring, this study analyzes the changes in principal stress direction and the characteristics of microseismic event clustering to determine whether the fracture propagates directionally along the cut surface. The method for analyzing the changes in principal stress direction involves selecting stress monitoring points at specific locations near the cut surface, extracting their stress states at different times, and calculating the magnitude and direction of the principal stress at each moment. By comparing the changes in principal stress direction before and after the cut surface is implemented and during the working face advancement, it is determined whether the stress field undergoes a deflection favorable to fracture along the cut surface. The method for analyzing the characteristics of microseismic event clustering involves defining a buffer zone in three-dimensional space centered on the designed cut surface. The buffer zone is defined as a spatial area within ±0.5 meters of the theoretical position of the designed cut surface. The spatial distribution of newly occurring microseismic events is statistically analyzed as the working face advances, and the proportion of microseismic events falling within this buffer zone is calculated. This proportion is called the buffer zone microseismic event ratio. The basis for determining whether the fracture propagates directionally along the cut surface is the establishment of a principal stress direction change criterion and a microseismic event clustering criterion. The criterion for the change of principal stress direction requires that the angle between the direction of the maximum principal stress and the normal of the cut surface decreases to below the principal stress angle threshold during the critical period. The critical period refers to the period from when the working face advances to about 20 meters from the cut surface until the cut surface is fully exposed. The principal stress angle threshold is set based on rock fracture mechanics theory and field test data. When this angle is less than 30 degrees, the fracture surface is in a favorable stress state dominated by tensile failure. Therefore, the principal stress angle threshold is set at 30 degrees. The minimum angle calculated during the critical period is compared with the principal stress angle threshold of 30 degrees. The criterion for the clustering of microseismic events requires that when the working face approaches the cut area, the proportion of microseismic events in the buffer zone exceeds the microseismic event proportion threshold, and the events show a dense clustering phenomenon in time. The microseismic event proportion threshold is set based on the statistical analysis of the spatial distribution of microseismic events in historical successful roof cutting cases. Statistics show that this proportion is generally higher than 60% when there is effective directional fracture. Therefore, the microseismic event proportion threshold is set at 60%. The actual proportion of microseismic events in the buffer zone is compared with the microseismic event proportion threshold of 60%. When both the principal stress direction change criterion and the microseismic event accumulation criterion are met, i.e. the minimum included angle is less than 30 degrees during the critical period and the actual proportion of microseismic events in the buffer zone is greater than 60%, it is determined that the fracture is directionally propagating along the cut surface; if only one criterion is met or neither is met, it is determined that the fracture propagation direction deviates from the design cut surface.

[0082] The effectiveness of the key causal link is evaluated by combining the judgment results of the cut surface forming quality and whether the fracture propagates directionally along the cut surface. The evaluation process first quantifies and assigns values ​​to the two independent judgment results. A high score of 1 is assigned to a cut surface forming quality assessment of excellent or good, and a low score of 0 is assigned to a forming quality assessment of medium or poor. A high score of 1 is assigned to a judgment that the fracture propagates directionally along the cut surface, and a low score of 0 is assigned to a judgment that does not exist. Then, a comprehensive effectiveness threshold is set for the final judgment. The comprehensive effectiveness threshold is determined based on statistical analysis of the combination of two scores in historical successful and failed cases. Statistical analysis shows that when both independent judgments are successful, the total score is 2, corresponding to a successful case; when only one is successful, the total score is 1, corresponding to an uncertain or failed case. Therefore, the comprehensive effectiveness threshold distinguishing success from failure is set to 1.5. During the evaluation, the sum of the two scores is compared with the comprehensive effectiveness threshold of 1.5. If the total score reaches or exceeds 1.5, the key causal link from cut surface implementation to fracture propagation is evaluated as effective; if the total score is below 1.5, the effectiveness of this link is evaluated as insufficient.

[0083] Monitoring the size and accumulation morphology of the collapsed roof behind the goaf is crucial for assessing whether effective support has been formed for the basic roof. Methods for monitoring the size of collapsed blocks include reserving observation windows on the retaining wall structure in the roadway, using high-definition cameras to periodically photograph or record the goaf, or using a laser scanner to perform a 3D scan of the collapsed rock surface. Image analysis or point cloud processing is used to identify the approximate size of individual rock blocks and statistically analyze the block size distribution across different size ranges, such as the percentage of rock blocks larger than 0.5 meters, between 0.2 and 0.5 meters, and smaller than 0.2 meters in volume. Monitoring the accumulation morphology primarily involves measuring the height of the rock pile (the height of the pile surface from the roof) and observing whether the pile is dense and free of voids. The assessment of whether effective support has been formed is based on setting thresholds for block size distribution and roof contact height. The block size distribution threshold requires that moderately sized rock blocks dominate. This threshold is determined based on the theory of granular media bearing capacity and laboratory simulations. When the volume percentage of rock blocks with a size between 0.2 meters and 0.5 meters is greater than 50%, its bearing capacity and compressive strength are superior. Therefore, the block size distribution threshold is set at 50%. During evaluation, the actual monitored volume percentage of rock blocks with a size between 0.2 meters and 0.5 meters is compared with the block size distribution threshold of 50%. The roof contact height threshold requires that the void height between the surface of the gangue pile and the roof is less than the critical roof contact height, and there are no obvious continuous cavities within the observation range. The critical roof contact height is set based on the engineering requirement of ensuring timely support for the overlying rock strata. It is usually required that this height be less than 0.3 meters. Therefore, the critical roof contact height threshold is set at 0.3 meters. During evaluation, the actual measured maximum void height between the surface of the gangue pile and the roof is compared with the critical roof contact height threshold of 0.3 meters. If the monitoring results simultaneously meet the block size distribution threshold and the top height threshold, that is, the actual moderate block size percentage is greater than 50% and the maximum gap height is less than 0.3 meters, then the assessment shows that effective support has been formed for the basic top; otherwise, the assessment shows that the support is insufficient.

[0084] The completeness of the process chain is determined by correlating the assessment results of roof collapse block size, stacking morphology, and cut surface forming quality. The method of correlation analysis is to establish a process chain event table, which sequentially records the events and states of key causal links, including the cut surface forming quality assessment result, fracture directional propagation judgment result, roof collapse block size assessment result, and stacking morphology assessment result. The logical consistency between these event states is analyzed. Theoretically, a cut surface forming quality assessment of "excellent" or "good" and a fracture directional propagation judgment that is valid should result in the roof collapse block size meeting the block size distribution threshold and the stacking morphology meeting the roof connection height threshold. If the actual monitoring shows that both block size and stacking morphology meet the threshold requirements, it indicates that the process chain is complete and the causal transmission is smooth. If the cut surface forming quality assessment of "excellent" or "good" and a fracture directional propagation judgment that is valid, but the monitored block size does not meet the block size distribution threshold or the stacking does not meet the roof connection height threshold, it indicates that the process chain is interrupted at the collapse or stacking stage. The criterion for judging the completeness of the process chain is to check for any such logical inconsistencies or deviations from theoretical expectations. Based on the degree of logical consistency of all links in the process chain, the integrity of the process chain is judged as complete, basically complete, or broken.

[0085] Based on the process chain integrity assessment, the effectiveness of the causal links throughout the entire process from cut implementation to filling and support is comprehensively evaluated. This comprehensive evaluation integrates the process chain integrity assessment results with the effectiveness assessment results from cut implementation to fracture propagation. A comprehensive process effectiveness rating rule is established. When the process chain integrity assessment is complete, and the cut implementation to fracture propagation assessment is effective, the comprehensive evaluation indicates high causal link effectiveness. When the process chain integrity assessment is basically complete, and the cut implementation to fracture propagation assessment is effective or at a critical state, the comprehensive evaluation indicates medium causal link effectiveness. When the process chain integrity assessment is broken, or the cut implementation to fracture propagation assessment is ineffective, the comprehensive evaluation indicates low causal link effectiveness. This comprehensive evaluation result serves as an integrated indicator to characterize whether the entire chain process was successfully implemented as designed.

[0086] After the lower coal seam working face is mined, based on the evaluation results of the effectiveness of the key causal links in the chain reaction process completed in step S4, the roof strata above the directional pre-splitting cut form the expected collapse pattern and collapse, forming broken and distended gangue. Based on the evaluation of the effectiveness of the key causal links, the synergistic state of fracture propagation and roof rotation is determined. The evaluation results of the effectiveness of the key causal links are derived from the output of step S4. Determining the synergistic state requires the introduction of real-time strata movement monitoring data; a deep baseline displacement monitoring system is pre-installed in the roof strata within a certain range above and behind the cut surface. This system is implemented by drilling holes at different depths in the roof and installing anchor points and rope-type displacement gauges, for example, monitoring points are set at depths of 2 meters, 4 meters, and 6 meters above the cut surface. The determination of the synergistic state depends on the time series analysis of two core parameters: the location of the fracture propagation front and the rotation angle of the key roof strata. The location of the fracture propagation front is continuously obtained through the microseismic event clustering characteristic analysis in step S4, and is represented by the length of the fracture front from the starting end of the cut surface. The rotation angle of the key roof strata is calculated using deep baseline displacement monitoring data. The calculation method uses the roof at the roadway coal face as a relatively fixed point, calculating the ratio of the subsidence of different depth baselines relative to this fixed point to the horizontal distance. The arctangent value of this ratio is the approximate rotation angle of that stratum. The criteria for judging the coordinated state are set as a time synchronization threshold and a displacement correlation threshold. The time synchronization threshold is set based on the analysis of observational data on the coordination of rock strata movement in historical mining. Observational data shows that when the time difference between fracture propagation and the start of roof rotation exceeds 2 hours, their movements are often disconnected. Therefore, the time synchronization threshold is set to 2 hours. The judgment is made by comparing the difference between the actual time the fracture front reaches a certain position and the time when the rotation angle of the key monitoring point above that position begins to increase significantly with the time synchronization threshold of 2 hours. The displacement correlation threshold is set based on statistical analysis theory. It is generally considered that a Pearson correlation coefficient greater than 0.8 indicates a high correlation between two variables; therefore, the displacement correlation threshold is set to 0.8. The judgment is made by calculating the Pearson correlation coefficient between the fracture front position sequence and the rotation angle sequence of key monitoring points, and comparing this coefficient with the displacement correlation threshold of 0.8. If the actual time difference is less than the time synchronization threshold of 2 hours and the calculated correlation coefficient is greater than the displacement correlation threshold of 0.8, then the fracture propagation and roof rotation are considered to be in a coordinated state. If the actual time difference is greater than or equal to the time synchronization threshold of 2 hours or the calculated correlation coefficient is less than or equal to the displacement correlation threshold of 0.8, then the coordination is considered to be disordered.

[0087] After the coal seam working face is mined, under the guiding and synergistic effect of directional pre-splitting cuts, the roof strata above the cuts fracture along the cut surface and collapse. Working face mining refers to the process where the coal mining machine cuts the coal body along the working face, and the hydraulic supports move forward, causing the roof in front of the cut to lose direct support. Physically, the guiding effect of directional pre-splitting cuts manifests as the cut surface acting as a continuous weak surface, with tensile and shear strengths far lower than intact rock mass, becoming the preferred path for stress concentration and fracture development. Under synergistic conditions, fracture propagation and roof rotation mutually promote each other; the fracture provides a fulcrum for rotation, and the tensile stress generated by rotation further drives the fracture forward. Specifically, the fracture of the roof strata along the cut surface is manifested as follows: under mine pressure, cracks initiate from stress concentration points at the ends or middle of the cut surface and strictly extend along the spatial geometric trajectory of the cut surface towards the depth of the roof and both sides of the strike, until the rock mass above the cut surface is completely severed from the coal wall behind it. In addition to relying on the microseismic monitoring in step S4, monitoring this process can also be achieved by pre-embedding a fiber optic sensing system within the cut surface. This system can directly depict the precise leading edge of fracture propagation by measuring abrupt changes in fiber strain distribution. The specific manifestation of roof rock strata rotational collapse is that the rock mass isolated by the fracture, under the influence of its own weight and the pressure of the overlying strata, rotates and sinks towards the goaf, with the fracture line as its hinge axis. This process can be clearly captured by a deep baseline displacement monitoring system, showing a sudden acceleration in the sinking velocity at each monitoring point, and the sinking curve exhibiting typical rotational motion characteristics, i.e., points far from the fracture line sink more, while points close to the fracture line sink less. The geometry of the collapsed rock mass formed in this stage is determined by the spatial orientation of the cut surface, the height of the cut, and the mechanical properties of the rock strata themselves.

[0088] Collapsed rock masses fracture within the goaf, expanding in volume to form boulders that fill the goaf. During their descent into the goaf and impact with the floor, the collapsed rock mass undergoes secondary fracturing due to the expansion of existing internal fissures and the collision and compression between rock fragments. The degree of fracturing is influenced by factors such as rock mass strength, joint development, and descent height. Volume expansion is an inevitable physical phenomenon that occurs when a rock mass transforms from a continuous state to a state of fragmented block accumulation; its degree is quantified by the boulders expansion coefficient. The boulders expansion coefficient is defined as the ratio of the accumulated volume of the fragmented, loose rock mass to the volume of the original intact rock mass. The formation of boulders and their filling of the goaf is a dynamic compaction process. In the initial stage, collapsed rock fragments naturally accumulate within the goaf, forming a loose mass with high porosity. At this point, the boulders expansion coefficient reaches its initial value, which can be obtained through field measurements or engineering analogies. For example, for medium-hard sandstone, engineering analogies show that the initial boulders expansion coefficient ranges from 1.3 to 1.5. As the working face continues to advance, the thickness of the overlying collapsed rock layer increases. Under the pressure of the overlying rock layer, the lower collapsed gangue is gradually compacted, reducing porosity and the coefficient of fragmentation, eventually tending towards a stable coefficient of fragmentation. The effectiveness assessment of the goaf filling is crucial to confirming whether effective support has been formed. In addition to the morphological monitoring already conducted in step S4, the assessment method also requires monitoring the stress recovery state within the goaf. Pressure cells are embedded within the gangue filling body to monitor the gradual establishment of vertical stress. When the filling body is compacted to a certain extent, sufficient to bear part of the load from the overlying rock layer, the pressure cell reading will show a steady increase and eventually tend towards a stable value. Criteria for effective filling include spatial and mechanical criteria. The spatial criterion requires that the gangue accumulation body is basically in contact with the roof, that is, the roof contact height threshold defined in step S4 is met, for example, the maximum void height is less than 0.3 meters. The mechanical criteria require that the stable vertical stress value established within the filling body is not lower than the minimum support stress threshold. The minimum support stress threshold is set based on a mechanical analysis balancing the load of the overlying strata with the stability of the roadway. This is achieved by calculating the expected proportion of the overlying strata load borne by the filling body, for example, setting the proportion to be no less than 20%, and then combining this with the average unit weight and height of the overlying strata to deduce the required minimum support stress threshold. Through this process, after the lower coal seam working face is mined, under effective roof-cutting guidance and coordinated movement mechanisms, the roof strata collapse, break, expand, and ultimately form a broken and expanded gangue filling body that can effectively support the overlying strata.

[0089] Based on the actual effects of backfilling the goaf with crushed gangue and the evaluation results of the effectiveness of key causal links in the chain reaction process, the roof cutting parameters are dynamically adjusted. The backfill density and roof contact condition after backfilling the goaf with crushed gangue are monitored to obtain the actual backfilling effect. Monitoring of backfill density is achieved by pre-embedding or later-stage drilling exploration boreholes within the gangue backfill behind the goaf, inserting borehole radar probes or resistivity logging probes into these boreholes. The borehole radar probe emits electromagnetic waves and receives reflected waves from interfaces with different dielectric constants within the backfill. By analyzing the travel time and amplitude of the reflected waves, the distribution and proportion of voids within the backfill are calculated, thus obtaining the average porosity as a quantitative indicator of backfill density. The resistivity logging probe measures the apparent resistivity of the backfill. Since the resistivity of air is much higher than that of rock fragments, the average porosity can also be estimated by establishing an empirical relationship between apparent resistivity and porosity. Monitoring of the roof contact condition was conducted using a laser rangefinder. A laser beam was emitted from inside the roadway towards the roof of the goaf and the surface of the gangue pile, measuring the distances at multiple measuring points. The vertical gap between the gangue pile surface and the roof was calculated using geometric relationships, and the average roof contact gap was statistically determined. The actual filling effect was characterized by two quantitative indicators: average porosity and average roof contact gap. Filling density thresholds and roof contact gap thresholds were set for effect evaluation. The filling density threshold was set based on laboratory tests of the bearing capacity of gangue granules under different compaction degrees. Test data showed that when the average porosity was below a critical value, the granules began to exhibit better bearing and force transmission characteristics. For example, the critical value was determined to be 25% through experimental analysis; therefore, the filling density threshold was set at 25%. The roof contact gap threshold was set based on the basic requirements for timely support in engineering. Typically, the average roof contact gap is required to be less than a critical value to ensure timely support; for example, this critical value was set at 0.3 meters; therefore, the roof contact gap threshold was set at 0.3 meters. To obtain the actual filling effect, the average porosity calculated by monitoring is compared with the filling density threshold of 25%, and the average gap between the top and bottom is compared with the gap threshold of 0.3 meters. If the average porosity is lower than the filling density threshold of 25% and the average gap between the top and bottom is less than the gap threshold of 0.3 meters, the actual filling effect is considered good. If the average porosity is higher than or equal to the filling density threshold of 25% or the average gap between the top and bottom is greater than or equal to the gap threshold of 0.3 meters, the actual filling effect is considered poor.

[0090] Combining the actual filling effect with the evaluation results of the effectiveness of key causal links, the weak links and design deviations in the cascading process are analyzed. The evaluation results of the effectiveness of key causal links are derived from the output of step S4, including the evaluation results of the cut surface forming quality, the judgment results of whether the fracture propagates directionally along the cut surface, and the comprehensive evaluation results of the effectiveness of all causal links in the entire process. A fault diagnosis association table is established during the analysis process. This table defines the possible weak links or design deviations corresponding to different combinations of key causal link effectiveness evaluation results and actual filling effects. The currently obtained key causal link effectiveness evaluation results and actual filling effects are matched with the entries in the fault diagnosis association table to determine the specific weak links or design deviation directions. The construction of the fault diagnosis association table is based on the mining strata control theory, historical case data, and logical reasoning from engineering experience. A diagnostic rule is defined as follows: if the cut surface forming quality evaluation is poor or medium, and the actual filling effect is poor, then the weak link is determined to be the cut surface forming link. Another diagnostic rule is defined as follows: If the quality assessment of the cut surface formation is excellent or good, but the judgment on whether the fracture propagates directionally along the cut surface is invalid, and the actual filling effect is poor, then the weak link is determined to be the fracture propagation link. A third diagnostic rule is defined as follows: If the quality assessment of the cut surface formation is excellent or good, the judgment on whether the fracture propagates directionally along the cut surface is valid, but the overall assessment of the effectiveness of all causal links in the entire process is moderate or low, and the actual filling effect is poor, then the weak link is determined to be the collapse and filling link. A fourth diagnostic rule is defined as follows: If the assessment results of the effectiveness of the key causal links are all good, i.e., the quality assessment of the cut surface formation is excellent or good, the judgment on fracture directional propagation is valid, and the overall effectiveness assessment is high, but the actual filling effect is poor, then a design deviation is determined to exist. Through matching the above rules, a clear diagnostic conclusion for the weak link or the direction of the design deviation is output.

[0091] Based on the analysis results, the cutting height or cutting position in the cutting parameters are corrected, and the corrected cutting parameters are used in the design and implementation of subsequent sections. If the analysis results determine that the weak link is the cut formation link or the fracture propagation link, and the main cause is local geological anomalies or construction errors, the correction measures focus on optimizing local construction techniques, such as adjusting blasting parameters. In this case, there is no need to correct the overall cutting design parameters. If the analysis results determine that the weak link is the collapse and filling link, or there is a clear design deviation in the cutting height, the cutting height is corrected. The principle of correcting the cutting height is to increase the volume of the collapsed rock mass to improve filling. The correction amount is determined by a combination of trial and error and theoretical calculation. Based on the actual fragmentation coefficient obtained in step S5, for example, if the actual fragmentation coefficient is 1.35, combined with the current designed cutting height, the actual collapse volume is calculated. The volume that the current goaf needs to be effectively filled is calculated. This volume is determined by the mining height, cutting height, collapse angle, and working face advance parameters. Set a target filling volume coefficient, which is the ratio of the desired volume of crushed and expanded gangue to the volume of the goaf to be filled. For example, set the target filling volume coefficient to 1.15 to ensure sufficient roof contact. Calculate the required roof cutting height to meet the target filling volume coefficient; this calculated value is the corrected suggested roof cutting height. If the calculated corrected roof cutting height exceeds the immediate roof range, it needs to be checked against the principles of bearing capacity characteristics of the immediate floor strata in step S2. If the analysis results indicate a design deviation in the roof cutting position, the roof cutting position is corrected. Correcting the roof cutting position requires re-analyzing the distribution characteristics of the non-uniform residual stress field obtained in step S1. In the section ahead of the roadway extension direction, re-identify the stress relief zone based on the latest stress distribution map. To ensure the corrected roof cutting position is located within the stress relief zone, and considering construction convenience, the coordinates of the roof cutting position are re-determined. The corrected roof cutting height or roof cutting position parameters are updated in the roof cutting parameter design table and used as design input for step S2 for the next mining cycle or the design of roof cutting parameters in adjacent unmined sections. The revised parameters are applied in the next directional pre-splitting cut.

[0092] Example 2: Figure 2 A schematic diagram of the dynamic design system for lower coal seam roof cutting parameters based on the fractured and infilled effect of the present invention is given. The dynamic design system for lower coal seam roof cutting parameters based on the fractured and infilled effect includes the following modules:

[0093] The stress assessment module is used to determine the distribution characteristics of the non-uniform residual stress field formed in the floor strata of the upper coal seam after mining, and to assess the structural bearing characteristics of the immediate floor strata of the lower coal seam.

[0094] The parameter design module is used to design the roof cutting parameters of the lower coal seam roadway based on the distribution characteristics of the non-uniform residual stress field and the structural bearing characteristics of the immediate floor strata.

[0095] The cutting module is used to implement directional pre-splitting cuts in the roof of the lower coal seam roadway according to the designed cutting parameters;

[0096] The process monitoring module is used to monitor and evaluate the effectiveness of key causal links in the chain reaction process from the implementation of directional pre-splitting cuts to the collapse of the roof strata;

[0097] The gangue generation module is used to, after the mining of the lower coal seam working face, based on the effectiveness of key causal links in the chain process, cause the roof strata above the directional pre-splitting cut to form the expected collapse pattern and collapse, forming crushed gangue.

[0098] The parameter adjustment module is used to dynamically adjust the roof cutting parameters based on the actual effect of the crushed gangue filling the goaf and the evaluation results of the effectiveness of key causal links in the chain process.

[0099] The calculations involved in the embodiments are all dimensionless numerical calculations, and the preset parameters and thresholds in the calculations are set by those skilled in the art according to the actual situation.

[0100] The above embodiments can be implemented, in whole or in part, by software, hardware, firmware, or any other combination thereof. When implemented using software, the above embodiments can be implemented, in whole or in part, in the form of a computer program product.

[0101] Those skilled in the art will recognize that the modules and algorithm steps of the various examples described in conjunction with the embodiments disclosed herein can be implemented in electronic hardware, or a combination of computer software and electronic hardware. Whether these functions are implemented in hardware or software depends on the specific application and inventive constraints of the technical solution. Those skilled in the art can use different methods to implement the described functions for each specific application, but such implementation should not be considered beyond the scope of this application.

[0102] In addition, the functional modules in the various embodiments of this application can be integrated into one processing module, or each module can exist physically separately, or two or more modules can be integrated into one module.

[0103] In the several embodiments provided in this application, it should be understood that the disclosed systems, apparatuses, and methods can be implemented in other ways. For example, the apparatus embodiments described above are merely illustrative; for instance, the division of modules is only a logical functional division, and in actual implementation, there may be other division methods. For example, multiple modules or components may be combined or integrated into another system, or some features may be ignored or not executed. Furthermore, the coupling or direct coupling or communication connection shown or discussed may be through some interfaces; the indirect coupling or communication connection between apparatuses or modules may be electrical, mechanical, or other forms.

[0104] The above description is merely a specific embodiment of this application, but the scope of protection of this application is not limited thereto. Any variations or substitutions that can be easily conceived by those skilled in the art within the scope of the technology disclosed in this application should be included within the scope of protection of this application. Therefore, the scope of protection of this application should be determined by the scope of the claims.

[0105] In conclusion, the above description is only a preferred embodiment of the present invention and is not intended to limit the present invention. Any modifications, equivalent substitutions, improvements, etc., made within the spirit and principles of the present invention should be included within the protection scope of the present invention.

Claims

1. A dynamic design method for roof cutting parameters of lower coal seams based on the effect of fragmentation and filling, characterized in that, Includes the following steps: S1: Determine the distribution characteristics of the non-uniform residual stress field formed in the floor strata of the upper coal seam after mining, and evaluate the structural bearing characteristics of the immediate floor strata of the lower coal seam. S2: Based on the distribution characteristics of the non-uniform residual stress field and the structural bearing characteristics of the direct floor strata, design the roof cutting parameters of the lower coal seam roadway. S3: Based on the designed roof cutting parameters, implement directional pre-splitting cuts in the roof of the lower coal seam roadway; S4: Monitor and evaluate the effectiveness of key causal links in the chain reaction from the implementation of directional pre-splitting cuts to the collapse of the roof strata, including: After directional pre-splitting, borehole television is used to detect the morphology of the cut surface and evaluate the quality of the cut surface formation. Microseismic and stress monitoring devices were deployed on both sides of the cut surface to collect spatiotemporal data on stress redistribution and micro-fracture events under the influence of mining. Based on spatiotemporal data, we analyze the changes in principal stress direction and the characteristics of microseismic event clustering to determine whether the fracture propagates directionally along the cut surface. The effectiveness of key causal links is evaluated by combining the results of the assessment of the quality of the cut surface formation and the direction of fracture propagation. When assessing the effectiveness of key causal links, the following are specifically included: Monitor the size and accumulation pattern of the collapsed roof behind the goaf to assess whether it forms effective support for the basic roof. The results of the assessment of the collapse block size, accumulation pattern and the forming quality of the cut surface of the top plate are correlated to determine the integrity of the process chain. Based on the judgment results of process chain integrity, the effectiveness of the causal links in the entire process from cut implementation to filling and support is comprehensively evaluated. S5: After the lower coal seam working face is mined, based on the effectiveness of the key causal links in the chain process, the roof strata above the directional pre-splitting cut form the expected collapse pattern and collapse, forming broken and swollen gangue. S6: Based on the actual effect of filling the goaf with crushed gangue and the evaluation results of the effectiveness of key causal links in the chain process, dynamically adjust the roof cutting parameters.

2. The dynamic design method for lower coal seam roof cutting parameters based on the fragmentation and filling effect according to claim 1, characterized in that, Determine the distribution characteristics of the non-uniform residual stress field formed in the floor strata after the mining of the upper coal seam, and assess the structural bearing characteristics of the immediate floor strata of the lower coal seam, including: In-situ stress data were obtained by drilling exploratory boreholes in the roof of the lower coal seam and measuring using the stress relief method. Inversion calculations were performed based on coal seam geological conditions and in-situ geostress data to determine the distribution characteristics of the non-uniform residual stress field. Mechanical parameters of the immediate floor strata were obtained by drilling core samples and conducting indoor rock mechanics tests. Based on mechanical parameters and the results of borehole inspection of cored holes, the structural bearing characteristics of the direct bottom strata are evaluated.

3. The dynamic design method for lower coal seam roof cutting parameters based on the fragmentation and filling effect according to claim 1, characterized in that, Based on the distribution characteristics of the non-uniform residual stress field and the structural bearing characteristics of the immediate floor strata, the roof cutting parameters of the lower coal seam roadway are designed, including: Based on the distribution characteristics of the non-uniform residual stress field, stress concentration regions and stress relief regions are identified, and the distribution of stress relief regions guides the determination of the cutting position. Based on the structural bearing characteristics of the direct bottom strata, the integrity and bearing capacity of the rock mass are determined, and the cutting height is determined based on the depth of the weak bearing capacity area. Based on the combined cutting location and cutting height, determine the cutting parameters to be used for implementation.

4. The dynamic design method for lower coal seam roof cutting parameters based on the fragmentation and filling effect according to claim 3, characterized in that, When determining the cut-off height based on the structural bearing characteristics of the immediate floor strata, the specific steps include: When the structural bearing characteristics of the direct bottom strata show the presence of weak interlayers or significant delamination, the design should be based on the principle of increasing the top cut height so that the collapsed rock mass can penetrate the weak layer. When the structural bearing characteristics of the direct bottom strata show that the rock mass is intact and has strong bearing capacity, the design should be based on the principle of keeping the cut-off height within the range of the direct top.

5. The dynamic design method for lower coal seam roof cutting parameters based on the fragmentation and filling effect according to claim 1, characterized in that, Based on the designed roof cutting parameters, directional pre-splitting cuts are implemented in the roof of the lower coal seam roadway, including: Based on the location of the cut-off roof, mark the drilling location at the corresponding position on the roof of the lower coal seam roadway; Determine the drilling depth and charge length based on the cut-off height; Drill holes to the predetermined depth at the marked location and fill them with shaped charge blasting devices. Then, carry out directional pre-splitting blasting along the design direction to form directional pre-splitting cuts.

6. The dynamic design method for lower coal seam roof cutting parameters based on the fragmentation and filling effect according to claim 1, characterized in that, After the lower coal seam working face is mined, based on the effectiveness of key causal links in the chain reaction process, the roof strata above the directional pre-splitting cuts form the expected collapse pattern and collapse, forming fragmented gangue, including: Based on the assessment of the effectiveness of key causal links, the synergistic state of fracture propagation and roof rotation is determined. After the coal seam working face is mined, under the guidance and coordination of directional pre-splitting cuts, the roof strata above the cuts fracture and collapse along the cut surface. The collapsed rock mass broke apart within the goaf, and its volume expanded to form boulders that filled the goaf.

7. The dynamic design method for lower coal seam roof cutting parameters based on the fragmentation and filling effect according to claim 1, characterized in that, Based on the actual effects of filling the goaf with crushed gangue and the evaluation results of the effectiveness of key causal links in the chain reaction process, the roof cutting parameters are dynamically adjusted, including: Monitor the filling density and roof connection of crushed gangue after filling the goaf to obtain the actual filling effect; Based on the evaluation results of actual filling effect and the effectiveness of key causal links, analyze the weak links and design deviations in the chain process; Based on the analysis results, the cutting height or cutting position in the cutting parameters are corrected, and the corrected cutting parameters are used in the design and implementation of subsequent sections.

8. A dynamic design system for lower coal seam roof cutting parameters based on the effect of fractured expansion and filling, used to implement the dynamic design method for lower coal seam roof cutting parameters based on the effect of fractured expansion and filling as described in any one of claims 1-7, characterized in that, Includes the following modules: The stress assessment module is used to determine the distribution characteristics of the non-uniform residual stress field formed in the floor strata of the upper coal seam after mining, and to assess the structural bearing characteristics of the immediate floor strata of the lower coal seam. The parameter design module is used to design the roof cutting parameters of the lower coal seam roadway based on the distribution characteristics of the non-uniform residual stress field and the structural bearing characteristics of the immediate floor strata. The cutting module is used to implement directional pre-splitting cuts in the roof of the lower coal seam roadway according to the designed cutting parameters; The process monitoring module is used to monitor and evaluate the effectiveness of key causal links in the chain reaction process from the implementation of directional pre-splitting cuts to the collapse of the roof strata; The gangue generation module is used to, after the mining of the lower coal seam working face, based on the effectiveness of key causal links in the chain process, cause the roof strata above the directional pre-splitting cut to form the expected collapse pattern and collapse, forming crushed gangue. The parameter adjustment module is used to dynamically adjust the roof cutting parameters based on the actual effect of the crushed gangue filling the goaf and the evaluation results of the effectiveness of key causal links in the chain process.

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