Methods for controlling fracturing of surrounding rock in main roadways during the final mining stage of closely spaced coal seam superimposed working faces

Abstract

This invention discloses a method for controlling the fracturing of the surrounding rock in the main roadway during the final mining stage of a near-coal seam superimposed working face, belonging to the field of mine pressure and strata control technology. First, through rock mechanics testing and key stratum discrimination theory, the parameters of the key strata of the upper and lower coal seams' hard roofs are determined. Then, based on the deformation characteristics of the lower coal seam main roadway, the offset distance of the upper coal seam's stop-mining line is optimized and adjusted so that the stop-mining line is outside the boundary of the supporting pressure influence. A fan-shaped borehole is drilled in the roof of the upper coal seam roadway, and constant-displacement high-pressure hydraulic fracturing is used to form a directional through-fracture zone, blocking the transmission of mining stress to the upper coal seam main roadway. Simultaneously, a cross-layer floor borehole is drilled, and high-displacement pulse fracturing is used to weaken the key strata of the lower coal seam's roof, eliminating the influence of high static pressure and reducing the superposition of mining stress between the upper and lower coal seams. This invention can simultaneously protect the upper and lower coal seam main roadways using the upper coal seam roadway, with controllable pressure relief, safe construction, and low resource loss, significantly improving the stability of roadways in deep, near-coal seam group mining.

Description

Technical Field

[0001] This invention relates to the field of mine pressure and strata control technology, specifically to a method for controlling the fracturing of the surrounding rock in the main roadway during the final mining stage of a working face with closely spaced coal seams. Background Technology

[0002] As mining depths increase, deep rock masses are subjected to complex mechanical environments characterized by high ground stress, intense mining activity, and nonlinear rheology. This results in roadway surrounding rock exhibiting large deformation, long deformation duration, and susceptibility to support structure failure. Under these conditions, the superposition effect of mining stress fields is significantly amplified in near-distance coal seam overlapping mining faces. Peak support pressures are high, and the impact range is wide. Furthermore, the high stress generated by the upper coal seam mining is transmitted through the floor, continuously disturbing the lower coal seam. Especially now that the current coal seam mining face has been completed, the lower coal seam main roadway is not only continuously affected by the mining stress generated by the upper coal seam mining but also by the continuous high static pressure transmitted from the hard cantilever beam structure above the lower coal seam goaf. The combined effect of these two factors makes the deformation of the main roadway difficult to control, seriously threatening the safe production of the mine.

[0003] To address the aforementioned issues, traditional methods for controlling surrounding rock in roadways mainly include: increasing the width of the protective coal pillar, strengthening the support, and employing pressure-relief blasting. However, increasing the width of the protective coal pillar can lead to significant coal resource losses, and the adjustment space for the pillar width is limited by the mine development layout; strengthening the support (such as anchor cables, U-shaped steel supports, etc.) can only passively resist surrounding rock deformation and cannot fundamentally reduce the mining stress acting on the main roadways; while pressure-relief blasting can release surrounding rock stress to some extent, it has drawbacks such as secondary damage to the roadway surrounding rock caused by blasting vibration, high construction safety risks, and uncontrollable pressure relief range.

[0004] Hydraulic fracturing technology is widely used in coal mine roof control due to its advantages such as good controllability, large pressure relief range, and minimal damage to surrounding rock. This technology uses high-pressure pumps to inject fracturing fluid to create fractures in the rock strata, altering the mechanical properties and integrity of the rock mass. This achieves the active control effect of reducing the overall strength of the rock mass and blocking the propagation path of mining stress. Currently, hydraulic fracturing technology has achieved significant results in areas such as treating hard roofs in coal mines, roadway stabilization with roof cutting for pressure relief, and preventing rockbursts, becoming an important means of controlling surrounding rock in deep mines.

[0005] Regarding the control of surrounding rock deformation in the main roadway of a coal seam superimposed working face, existing hydraulic fracturing methods mostly focus on the decompression protection of a single coal seam or a single roadway, lacking a systematic technical solution for the coordinated protection of the main roadways of the upper and lower coal seams. When the coal seam has been mined out and the roadway is closed and nearby construction is not possible, there is a lack of effective active decompression means. Existing technologies fail to fully utilize the upper coal seam roadway as a construction channel, while simultaneously decompressing the roof of the upper coal seam main roadway and eliminating the impact of the goaf roof in the lower coal seam main roadway to achieve coordinated protection.

[0006] Therefore, there is an urgent need to develop a pressure relief method suitable for working faces with closely spaced coal seams, which can simultaneously control the deformation of the surrounding rock in the main roadways of the upper and lower coal seams using the upper coal seam roadway, so as to achieve long-term stability of the roadway surrounding rock under the conditions of close-spaced coal seam mining in deep mines. Summary of the Invention

[0007] The purpose of this invention is to provide a method for controlling the fracturing of the surrounding rock in the main roadway during the final mining period of a working face with closely spaced coal seams. This method addresses the challenge of controlling the deformation of the surrounding rock in the main roadway during the final mining period of a working face with closely spaced coal seams, reduces the impact of working face mining on the main roadway, and addresses the problem of continuous deformation in the main roadway corresponding to a working face that has been mined out. By utilizing the upper coal seam roadway to simultaneously control the deformation of the surrounding rock in the main roadways of both upper and lower coal seams, the method achieves long-term stability of the surrounding rock in the roadway under the conditions of mining closely spaced coal seams in deep mines.

[0008] To achieve the above objectives, the present invention is conceived as follows: by employing hydraulic fracturing technology, the dynamic stress source of mining is reduced by fracturing the basic roof of the upper coal seam, the dynamic stress transmission path of the upper coal seam is blocked by fracturing the basic bottom, and the static stress source is reduced by fracturing the hard overhang of the lower coal seam. This reduces the impact of the working face mining on the main roadway of the nearby coal seam group mining area.

[0009] The technical solution adopted in this invention is: a method for controlling the fracturing of the surrounding rock in the main roadway during the final mining period of a closely spaced coal seam superimposed working face, comprising the following steps: Step S1, conducting rock mechanics tests on the roof strata of the upper coal seam and the interlayer strata between the upper and lower coal seams; using the key layer discrimination theory to determine the position and thickness of the low-level hard roof key layer and the high-level hard roof key layer above the upper coal seam, and to determine the position and thickness of the hard roof key layer above the lower coal seam; Step S2, based on the deformation characteristics of the main roadway of the lower coal seam, combined with the spatial superposition relationship of the upper and lower coal seam working faces and the superposition of mining stress, determining the spatial position of the upper coal seam stop mining line relative to the lower coal seam stop mining line, establishing a mechanical equilibrium model of the interlayer rock pillar, and calculating the optimal offset distance of the upper coal seam stop mining line relative to the lower coal seam stop mining line.

[0010] Step S3: In the track roadway and transport roadway of the upper coal seam working face, according to the position and thickness of the key layers of the low and high hard roof determined in step S1, and the position of the upper coal seam stop line determined in step S2, construct fan-shaped boreholes in the roof of the upper coal seam behind the stop line to perform hydraulic fracturing on the key layers of the low and high hard roof, forming a directional through-overburden fracture zone dominated by a single main fracture.

[0011] Step S4: In the track roadway and transport roadway of the upper coal seam working face, construct a fan-shaped borehole that penetrates the interlayer rock strata. The borehole ends in the key layer of the hard roof above the lower coal seam. Use high-volume pulse fracturing to hydraulically fracture the key layer of the hard roof above the lower coal seam, forming a dense fracture network. This allows the hard roof above the lower coal seam to fully collapse. The collapsing gangue is used to fill the goaf to support the roof and prevent the main roadway of the lower coal seam from being continuously affected by the high static pressure of the hard roof in the goaf.

[0012] Step S5: Set up roadway deformation monitoring points in the upper and lower coal seam main roadways to monitor the roadway convergence, floor heave, and anchor cable stress changes. When the monitoring data shows that the roadway deformation rate tends to stabilize and the cumulative deformation is less than the safety set value, it is determined that the water pressure fracturing control of roadway deformation has met the standard.

[0013] As a further improvement of the present invention, in step S1, the rock mechanics test specifically involves drilling rock cores for indoor physical and mechanical tests to determine the uniaxial compressive strength, tensile strength, elastic modulus, Poisson's ratio, and internal friction angle of each rock layer.

[0014] As a further improvement of the present invention, in step S2, a mechanical equilibrium model of the interlayer rock column is established, including the following steps: Step 1, determine the mechanical calculation object: the rock strata between the upper coal seam stop line and the lower coal seam stop line, and located between the upper coal seam and the lower coal seam, are determined as the interlayer rock column, and the horizontal distance between the upper coal seam stop line and the lower coal seam stop line is used as a spatial parameter characterizing the stress range of the interlayer rock column.

[0015] The second step is to establish an equivalent mechanical structure: the hard roof structure that has not fully collapsed between the edge of the goaf and the stop line of the upper coal seam after mining is equivalent to a cantilever bearing structure that bears the load of the overlying strata, and the interlayer rock pillars are equivalent to a continuous rock mass structure that can transmit the mining stress of the upper coal seam.

[0016] The third step is to determine the stress transfer relationship: taking the cantilever bearing structure near the coal seam cessation line as the source of the upper load, we analyze the load transfer effect of its bending and sinking and end stress on the upper interface of the interlayer rock column, and determine the constraint relationship between the vertical stress and the horizontal stress in the interlayer rock column.

[0017] Step 4: Establish the equilibrium relationship between inter-layer rock pillars: Based on the load transmitted from the hard roof structure of the upper coal seam to the inter-layer rock pillars, the lateral constraint effect of the inter-layer rock pillars themselves, and the stress continuity at the interface between the upper and lower coal seams, determine the correspondence between the stress distribution within the inter-layer rock pillars and the relative positions of the stop mining lines of the upper and lower coal seams.

[0018] Step 5: Establish the criteria for the instability of the surrounding rock in the lower coal seam main roadway: Based on the measured deformation characteristics of the lower coal seam main roadway, the horizontal stress corresponding to the shrinkage deformation of the two sides and the vertical stress corresponding to the heave deformation of the bottom are taken as the main control factors, and the surrounding rock strength conditions are combined to determine whether the surrounding rock in the lower coal seam main roadway is in a stable state.

[0019] Step 6: Determine the basis for calculating the offset of the stop mining line: Based on the inter-layer rock pillar balance relationship, stress transmission relationship and the instability criterion of the surrounding rock of the lower coal seam main roadway, determine the calculation basis for adjusting the stop mining line of the upper coal seam to move inward, outward or slightly offset relative to the stop mining line of the lower coal seam.

[0020] As a further improvement of the present invention, in step S2, the calculation of the optimal offset distance between the upper coal seam stop line and the lower coal seam stop line is specifically as follows: Step S21, when the deformation of the lower coal seam main roadway is mainly due to the contraction of the two sides, an inward movement mode is adopted, that is, the upper coal seam stop line moves towards the goaf, and the upper coal seam stop line is located inside the lower coal seam stop line. The optimal offset distance ΔL1 is calculated according to the following formula: In the formula, ΔL1 is the optimal distance (m) between the upper coal seam stop line and the lower coal seam stop line; L u σ is the length of the fixed-support cantilever beam in the upper coal seam, in meters; h,min The shear strength of the surrounding rock in the lower coal seam main roadway is given in MPa; L p φ is the width of the interlayer rock pillar, in meters; k is the horizontal distance between the stop mining lines of the upper and lower coal seams; q is the lateral pressure coefficient; q is the load of the overlying strata, in MPa; and φ is the internal friction angle of the interlayer rock pillar.

[0021] Step S22: When the deformation of the main roadway of the lower coal seam is mainly characterized by floor heave, the outward shift mode is adopted, that is, the stop line of the upper coal seam is moved towards the main roadway, and the stop line of the upper coal seam is located outside the stop line of the lower coal seam. The optimal offset distance ΔL2 is calculated by the following formula: In the formula, ΔL2 is the optimal distance (m) between the upper coal seam stop line and the lower coal seam stop line; σ v,min denoted as σ, where σ is the compressive strength of the surrounding rock in the main roadway of the lower coal seam, expressed in MPa.

[0022] Step S23: When the deformation of the main roadway of the current coal seam is accompanied by both side shrinkage and floor heave, the alignment mode or small offset mode is adopted, and the offset amount ΔL3 is calculated according to the following formula: In the formula, α is a weighting coefficient, which is determined based on the ratio of the shrinkage of the two sides to the deformation of the bottom drum measured on site, and the value range is 0~1.

[0023] The offset distance ΔL is the horizontal movement distance of the upper coal seam stop line relative to its initial position. When the inward movement mode is adopted, ΔL=ΔL1; when the outward movement mode is adopted, ΔL=ΔL2; when the small offset mode is adopted, ΔL=ΔL3; when ΔL>0, it means that the upper coal seam stop line moves towards the goaf; when ΔL<0, it means that the upper coal seam stop line moves towards the main roadway.

[0024] The initial width Lp of the inter-layer rock pillar before the adjustment of the coal seam stop line is used as the reference width. The width of the inter-layer rock pillar is then optimized and adjusted according to the offset distance ΔL to obtain the adjusted width of the inter-layer rock pillar. It is calculated using the following formula: Based on the above calculation results, the stop mining line of the upper coal seam is determined to be located inside, outside or aligned with the stop mining line of the lower coal seam, ensuring that the re-determined stop mining line is located outside or at the boundary of the support pressure influence.

[0025] As a further improvement of the present invention, in step S23, the weighting coefficient α is based on the on-site measured deformation ε of the two sides. h With the deformation amount ε of the bottom drum v The ratio is determined as follows: When α≥0.7, it is determined that the deformation is mainly due to contraction of the two sides.

[0026] When α≤0.3, it is determined that the bottom drum is the main component.

[0027] When 0.3 < α < 0.7, it is determined to be a composite type.

[0028] As a further improvement of the present invention, in step S3, the opening position of the top plate drilling hole is located within 5-10m behind the stop mining line.

[0029] As a further improvement of the present invention, in step S3, the drilling and fracturing work in the upper coal seam working face track roadway and the upper coal seam working face transport roadway is completed when the working face advances to a distance of not less than 200m from the predetermined stop line.

[0030] As a further improvement of the present invention, in step S3, the spacing between the final holes of the fan-shaped upper coal seam roof boreholes is 15-20m, and the line connecting the final hole positions is parallel to the working face arrangement direction.

[0031] As a further improvement of the present invention, in step S5, the safety setting is: the convergence rate of the two sides of the roadway is less than 1 mm / d, the floor heave rate is less than 0.5 mm / d, the cumulative convergence amount is less than 80% of the design allowable value, and the force fluctuation amplitude of the anchor cable is less than 30% of its yield load.

[0032] Compared with the prior art, the present invention has the following beneficial effects: 1. The present invention reduces the integrity, bending stiffness and fracture energy of the key layers of the hard roof at low and high positions of the upper coal seam by implementing constant discharge water pressure fracturing, thereby weakening the mining dynamic pressure source generated by roof fracture during the final mining period; at the same time, it forms a directional through fracture zone, blocking or attenuating the propagation path of mining stress to the main roadway of the upper coal seam, thereby achieving active pressure relief protection of the main roadway of the upper coal seam.

[0033] 2. This invention utilizes the downward drilling of the upper coal seam through the bottom plate to perform high-volume pulse fracturing on the key layer of the hard roof above the goaf of the lower coal seam, forming a dense fracture network, weakening the bearing capacity of the hard overhanging roof and promoting its full collapse. The collapsed gangue is then used to fill the goaf, eliminating or reducing the continuous high static pressure effect of the hard overhanging roof on the main roadway of the lower coal seam from the source. It is especially suitable for situations where the lower coal seam has been mined out, the roadway is closed, or there are no nearby construction conditions.

[0034] 3. Based on the shrinkage, floor heave, or combined deformation characteristics of the two sides of the main roadway of the lower coal seam, this invention calculates and optimizes the offset distance of the stop line of the upper coal seam relative to the stop line of the lower coal seam, so that the peak value of the unmined support pressure of the upper coal seam and the weakened zone of the surrounding rock of the main roadway of the lower coal seam are spatially misaligned, avoiding the superposition of dynamic pressure, static pressure and support pressure peak values ​​in the surrounding rock of the main roadway, thereby reducing the combined disturbance impact under the condition of overlapping mining of closely spaced coal seams.

[0035] 4. This invention organically combines the optimization of the stop-mining line offset, the fracturing of the upper coal seam roof, the fracturing of the lower coal seam roof, and the monitoring of roadway deformation to construct a closed-loop control system of "source weakening - path blocking - spatial avoidance - dynamic verification". Compared with increasing the coal pillar, simply strengthening the support, or a single fracturing measure, it can reduce coal resource loss and construction risks while achieving coordinated, quantifiable, and long-term stable control of the deformation of the surrounding rock in the main roadways of the upper and lower coal seams. Attached Figure Description

[0036] Figure 1 It is a coal mining scenario layout diagram on which the method of controlling the fracturing of the surrounding rock in the main roadway during the final mining stage of a working face with closely spaced coal seams is based.

[0037] Figure 2 yes Figure 1 Cross-sectional view along the advancing direction of the working face.

[0038] Explanation of reference numerals in the attached diagram: 1-Upper coal seam; 2-Key stratum of the lower hard roof of the upper coal seam; 3-Key stratum of the upper hard roof of the upper coal seam; 4-Lower coal seam; 5-Key stratum of the hard roof of the lower coal seam; 6-Location of the stop line of the upper coal seam working face; 7-Traction roadway of the upper coal seam working face; 8-Transport roadway of the upper coal seam working face; 9-Fan-shaped borehole in the roof of the upper coal seam; 10-Overlying fracture zone; 11-Main roadway of the upper coal seam; 12-Fan-shaped borehole in the floor of the cross-layer; 13-Dense fracture network; 14-Main roadway of the lower coal seam; 15-Hard overhang in the goaf of the lower coal seam; 16-Goaf of the upper coal seam; 17-Upper coal seam longwall face; 18-Shaft; 19-Location of the stop line of the lower coal seam working face. Detailed Implementation

[0039] To make the objectives, technical solutions, and advantages of the embodiments of the present invention clearer, the technical solutions of the embodiments of the present invention will be clearly and completely described below. Obviously, the described embodiments are only some, not all, of the embodiments of the present invention. All other embodiments obtained by those skilled in the art based on the described embodiments of the present invention are within the scope of protection of the present invention.

[0040] Example 1

[0041] Please refer to Figure 1 and Figure 2 This invention proposes a method for controlling the fracturing of the surrounding rock in the main roadway during the final mining stage of a coal seam superimposed working face with closely spaced coal seams, comprising the following steps: Step S1, Key Layer Identification and Mechanical Testing: Rock mechanics testing is conducted on the roof strata of the upper coal seam and the interlayer strata between the upper coal seam 1 and the lower coal seam 4, including drilling rock cores for indoor physical and mechanical tests to determine the uniaxial compressive strength, tensile strength, elastic modulus, Poisson's ratio, and internal friction angle of each stratum. The key layer identification theory is used to determine the position and thickness of the key layer 2 (lower hard roof strata of the upper coal seam above 1) and the key layer 3 (higher hard roof strata of the upper coal seam above 1), as well as the position and thickness of the key layer 5 (hard roof strata of the lower coal seam above 4). The key layer identification theory is as follows: Key layers are characterized by greater thickness and higher strength compared to other strata. Their fracture can affect the synchronous fracture of the local overburden field. They are both good load-bearing structures and highly hazardous load-bearing structures. The identification criteria are as follows: In the formula, q n+1 q n These represent the loads on the first critical layer at the (n+1)th and nth layers, respectively. The process proceeds layer by layer from bottom to top until the location of the uppermost potentially critical hard rock layer is determined, assuming that k layers of overburden satisfy the requirements of this formula. (j=1,2,···,k); where l j+1 , l jThese are the fracture distances of the (j+1)th and (j)th rock layers, respectively. If the hard rock layer j does not satisfy this formula, then all the rock layer loads controlled by the (j+1)th hard rock layer should be applied to the kth layer, and the fracture distance of the hard rock layer k should be recalculated before continuing the judgment.

[0042] Specifically, the key layer 2 of the low-level hard roof of the upper coal seam is the thick hard rock layer closest to the upper coal seam 1, and the key layer 3 of the high-level hard roof of the upper coal seam is a hard rock layer that is located above the key layer 2 of the low-level hard roof of the upper coal seam and is thicker or stronger.

[0043] exist Figure 1 In the middle, behind the upper coal seam mining face 17 is the upper coal seam goaf 16, and in front is the upper coal seam working face stop line position 6. The shaft 18 is drilled from the ground to the lower coal seam main roadway 14. The lower coal seam main roadway 14 is parallel to the upper coal seam main roadway 11, and both are perpendicular to and connected to the shaft 18.

[0044] Step S2, Re-determining the Stop-Mining Line and Optimizing its Spatial Relationship: Given that the mining of the lower coal seam 4 has been completed and the main roadway 14 of the lower coal seam continues to deform after the roadway is closed, the stop-mining line position 6 of the upper coal seam working face is re-determined.

[0045] Based on the deformation characteristics of the lower coal seam roadway 14, and considering the spatial superposition relationship between the upper and lower coal seam working faces and the superposition of mining-induced stresses, the relative spatial relationship between the stop-mining line position 6 of the upper coal seam working face and the stop-mining line position 19 of the lower coal seam working face is determined, and a mechanical equilibrium model of the interlayer rock pillar is established. Specifically, the interlayer rock pillar refers to the rock strata located between the upper and lower coal seams, between the stop-mining lines of the upper and lower coal seams. Under the condition of spatial superposition of closely spaced coal seam groups, after the upper coal seam is mined, a cantilever beam structure is formed at the edge of the goaf. This cantilever beam transfers the load of the overlying strata to the interlayer rock pillar, and then from the interlayer rock pillar to the lower coal seam roadway 14.

[0046] To quantitatively describe this stress transfer mechanism, the following mechanical equilibrium model is established: First, determine the mechanical calculation object: the rock strata between the upper coal seam stop line and the lower coal seam stop line, and located between the upper coal seam and the lower coal seam, are defined as interlayer rock pillars, and the horizontal distance between the upper coal seam stop line and the lower coal seam stop line is used as a spatial parameter characterizing the stress range of the interlayer rock pillar.

[0047] The second step is to establish an equivalent mechanical structure: the hard roof structure that has not fully collapsed between the edge of the goaf and the stop line of the upper coal seam after mining is equivalent to a cantilever bearing structure that bears the load of the overlying strata, and the interlayer rock pillars are equivalent to a continuous rock mass structure that can transmit the mining stress of the upper coal seam.

[0048] The third step is to determine the stress transfer relationship: taking the cantilever bearing structure near the coal seam cessation line as the source of the upper load, we analyze the load transfer effect of its bending and sinking and end stress on the upper interface of the interlayer rock column, and determine the constraint relationship between the vertical stress and the horizontal stress in the interlayer rock column.

[0049] Step 4: Establish the equilibrium relationship between inter-layer rock pillars: Based on the load transmitted from the hard roof structure of the upper coal seam to the inter-layer rock pillars, the lateral constraint effect of the inter-layer rock pillars themselves, and the stress continuity at the interface between the upper and lower coal seams, determine the correspondence between the stress distribution within the inter-layer rock pillars and the relative positions of the stop mining lines of the upper and lower coal seams.

[0050] Step 5: Establish the criteria for the instability of the surrounding rock in the lower coal seam main roadway: Based on the measured deformation characteristics of the lower coal seam main roadway, the horizontal stress corresponding to the shrinkage deformation of the two sides and the vertical stress corresponding to the heave deformation of the bottom are taken as the main control factors, and the surrounding rock strength conditions are combined to determine whether the surrounding rock in the lower coal seam main roadway is in a stable state.

[0051] Step 6: Determine the basis for calculating the offset of the stop mining line: Based on the inter-layer rock pillar balance relationship, stress transmission relationship and the instability criterion of the surrounding rock of the lower coal seam main roadway, determine the calculation basis for adjusting the stop mining line of the upper coal seam to move inward, outward or slightly offset relative to the stop mining line of the lower coal seam.

[0052] The optimal offset distance between the upper coal seam stop line and the lower coal seam stop line is calculated according to the following theoretical formula: Step S21: When the deformation of the lower coal seam main roadway 14 is mainly due to the contraction of the two sides, the inward shift mode is adopted. The inward shift mode is suitable for the two sides contraction as the main factor, in order to reduce the concentration of horizontal stress. Specifically, the stop line position 6 of the upper coal seam working face is located inside the stop line position 19 of the lower coal seam working face, and the optimal offset distance ΔL1 is calculated according to the following formula: In the formula, ΔL1 is the optimal distance (m) between the upper coal seam stop line and the lower coal seam stop line; L u σ is the length of the fixed-support cantilever beam in the upper coal seam, in meters; h,min The shear strength (MPa) of the surrounding rock in the lower coal seam main roadway; L p φ is the width of the interlayer rock pillar, in meters, which is the horizontal distance between the stop mining lines of the upper and lower coal seams; k is the lateral pressure coefficient; q is the load of the overlying strata, in MPa; and φ is the internal friction angle of the interlayer rock pillar.

[0053] Step S22: When the deformation of the main roadway 14 in the lower coal seam is mainly characterized by floor heave, the outward shift mode is adopted. The outward shift mode is suitable for floor heave as the main cause, in order to adjust the vertical load and stress transfer of the floor. Specifically, the stop line position 6 of the upper coal seam working face is located outside the stop line position 19 of the lower coal seam working face, and the optimal offset distance ΔL2 is calculated by the following formula: In the formula, ΔL2 is the optimal distance (m) between the upper coal seam stop line and the lower coal seam stop line; σ v,mindenoted as , which is the compressive strength of the surrounding rock in the main roadway of the lower coal seam, expressed in MPa.

[0054] Step S23: When the deformation of the main roadway in the current coal seam is characterized by both sides contracting and floor heave, an alignment mode or a small offset mode is adopted. The alignment / small offset mode is suitable for complex deformation and controls both horizontal and vertical stresses. Specifically, the offset ΔL3 is calculated using the following formula: In the formula, α is a weighting coefficient, which is determined based on the ratio of the shrinkage of the two sides to the deformation of the bottom drum measured on site, and the value range is 0~1.

[0055] Specifically, the weighting coefficient α is based on the on-site measured deformation ε of the two sides. h With the deformation amount ε of the bottom drum v The ratio is determined as follows: .

[0056] When α≥0.7, it is determined that the deformation is mainly due to shrinkage of the two sides; when α≤0.3, it is determined that the deformation is mainly due to bulging of the bottom; when 0.3<α<0.7, it is determined that it is a composite type.

[0057] The offset distance ΔL is the horizontal movement distance of the upper coal seam stop line relative to its initial position. Specifically, when using the inward movement mode, ΔL = ΔL1; when using the outward movement mode, ΔL = ΔL2; and when using the small offset mode, ΔL = ΔL3. When ΔL > 0, it indicates that the upper coal seam stop line moves towards the goaf (inward movement mode); when ΔL < 0, it indicates that the upper coal seam stop line moves towards the main roadway (outward movement mode).

[0058] The initial width Lp of the inter-layer rock pillar before the adjustment of the coal seam stop line is used as the reference width. The width of the inter-layer rock pillar is then optimized and adjusted according to the offset distance ΔL to obtain the adjusted width of the inter-layer rock pillar. ΔL; where, Optimized interlayer column width, in meters.

[0059] Based on the above calculation results, the stop-mining line position 6 of the upper coal seam working face is determined to be inside, outside or aligned with the stop-mining line position 19 of the lower coal seam working face, ensuring that the re-determined stop-mining line is located outside or at the boundary of the support pressure influence boundary.

[0060] In this step, the length L of the cantilever beam of the upper coal seam 1 is... u Specifically, this refers to the length of the exposed, uncollapsed, hard roof above the goaf after the upper coal seam has been mined, due to incomplete roof support from the goaf gangue or the high strength of the key strata. In close-range coal seam mining, the roof behind the upper coal seam goaf gradually collapses and compacts. The mining-induced stress acting on the lower coal seam's main roadway primarily originates from the exposed portion of the key strata, supported by the unmined coal mass in front of the working face, which is in a bent and subsiding state. Therefore, the equivalent fixed-support cantilever beam length in the mechanical model... The step size should be approximately equal to the period of the key layer.

[0061] Based on field experience in mine pressure observation, the periodic pressure step distance is usually 1 / 2 to 1 / 3 of the initial fracture distance.

[0062] In the formula, The step distance is determined by the period of this key layer, which is the length of the fixed cantilever beam, in meters. This is an empirical coefficient, typically taken as 0.3 to 0.5; The initial fracture distance of this critical layer is m.

[0063] In the formula, The load acting on this critical layer is measured in MPa. The thickness of this critical layer is m; The tensile strength of this critical layer is given in MPa.

[0064] Step S3, Constant Displacement Hydraulic Fracturing of the Roof: Within the track roadway 7 and transport roadway 8 of the upper coal seam working face, according to the positions and thicknesses of the key layers of the low and high hard roof determined in Step S1, and the positions of the stop-mining line redefined in Step S2, fan-shaped boreholes 9 are drilled in the roof of the upper coal seam behind the stop-mining line. A constant displacement high-pressure pump is used to hydraulically fracture the key layers of the low and high hard roof, maintaining a stable pumping pressure. This forms a directional, penetrating overlying fracture zone 10 dominated by a single main fracture, used to block the propagation path of mining stress towards the main roadway 11 of the upper coal seam and control the deformation of the main roadway 11. The so-called directional, penetrating overlying fracture zone 10 refers to hydraulic fracturing of the same hard rock layer in the roof. This type of rock layer possesses inherent properties such as hard texture, dense structure, and low porosity. Under constant displacement fracturing mode, fracture propagation always extends along the direction of maximum principal stress, and the overall direction of in-situ stress within the study area is consistent, thereby achieving the formation of directional and unified through fractures between adjacent boreholes. In addition, the hydraulically generated fractures from the early fracturing stage can guide and induce fractures in adjacent boreholes, further promoting the interconnection of fractures in each borehole.

[0065] The opening location of the roof boreholes is within 5-10m behind the stop line. The fracturing time for the boreholes in the upper coal seam working face track roadway 7 and transport roadway 8 must be 200m before the working face reaches the nearest main roadway in the advancing direction. For the fan-shaped arrangement of the upper coal seam roof boreholes, the elevation angle is determined based on the length of the upper coal seam working face and the thickness of the interlayer strata. The final borehole spacing is 15-20m, and the line connecting the final borehole locations is parallel to the working face layout direction. Specifically, hydraulic fracturing adopts a segmented retreating process. After fracturing the hard rock strata deep within the borehole, the process retreats to the next segment of hard rock strata to continue fracturing until the fracturing operation of the entire borehole is completed.

[0066] Step S4, High-volume pulse fracturing of the bottom plate: In the track roadway 7 and transport roadway 8 of the upper coal seam working face, a fan-shaped bottom plate borehole 12 is drilled downwards, penetrating the interlayer rock strata. The final borehole position is located in the key layer 5 of the hard roof of the lower coal seam above the lower coal seam 4. High-volume pulse fracturing is used to hydraulically fracture the key layer of the hard roof above the lower coal seam, forming a dense fracture network 13, weakening the bearing capacity of the hard rock strata, allowing the hard roof above the lower coal seam 4 to fully collapse. The crushing and swelling properties of the collapsed gangue are used to fill the goaf to support the roof, solving the continuous impact of the high static pressure of the hard suspended roof 15 of the lower coal seam goaf on the lower coal seam main roadway 14. In addition, the dense fracture network blocks the propagation path of the supporting pressure in the bottom plate, controlling the deformation of the lower coal seam main roadway 14. At the same time, after the collapse of the suspended roof of the lower coal seam goaf, the superposition of mining stress between the upper and lower coal seams can be reduced, further reducing the impact on the upper coal seam main roadway 11. Specifically, the drilling locations in the floor section are situated within the upper coal seam mining roadway, within a range of 5-10 meters behind the lower coal seam stop line. The drilling angle is determined based on the length of the lower coal seam working face and the thickness of the interlayer strata, ensuring that the drilling trajectory accurately enters the critical layer of the hard roof above the lower coal seam. The final hole spacing between each borehole is 15-20 meters, and the line connecting the final hole locations is parallel to the layout direction of the lower coal seam working face. When the real-time water pressure of the water pressure monitoring sensor remains below 5 MPa, or when water seeps from adjacent boreholes or the roof for more than 8 minutes, the high-pressure pump is shut off, the pressure relief valve connected to the high-pressure hose is opened to depressurize the pipeline, and the hydraulic fracturing is terminated.

[0067] Specifically, weakening the bearing capacity of hard rock strata includes weakening cantilever beam structures or weakening the rock strata as a whole. Weakening cantilever beam structures refers to destroying the integrity of the cantilever beam structure above the goaf of the lower coal seam through fracturing, cutting off high static pressure at the source; weakening the rock strata as a whole refers to forming a dense network of fractures in the target rock strata through fracturing, reducing the overall stiffness of the rock mass and blocking the stress propagation path.

[0068] Step S5, Effect Verification: Set up roadway deformation monitoring points in the upper and lower coal seam main roadways to continuously monitor the roadway convergence, floor heave, and anchor cable stress changes. When the monitoring data shows that the roadway deformation rate tends to stabilize and the cumulative deformation is less than the safety set value, it is determined that the water pressure fracturing control of roadway deformation has met the standard.

[0069] The safety setpoints are: the convergence rate of the two sides of the roadway is less than 1 mm / d, the floor heave rate is less than 0.5 mm / d, the cumulative convergence amount is less than 80% of the design allowable value, and the stress fluctuation amplitude of the anchor cable is less than 30% of its yield load.

[0070] This invention addresses the unique working condition during the final mining phase of closely spaced coal seams where both the upper and lower coal seam roadways are simultaneously subjected to mining-induced dynamic pressure and goaf static pressure. It establishes a collaborative control system of "source weakening—path blocking—spatial avoidance—dynamic verification." Specifically, during the final mining phase of the upper coal seam face, the periodic fracturing and subsidence of the key roof strata create significant mining-induced dynamic pressure sources, which are transmitted to the upper and lower coal seam roadways via overlying strata and interlayer pillars. To address this, this invention first constructs fan-shaped boreholes in the upper coal seam roadway to hydraulically fracture the low-lying and high-lying hard key roof strata. This fractures the originally intact, continuous, and high-bearing-capacity key strata, reducing their overall bending stiffness and fracture energy, and mitigating the dynamic load impact caused by sudden roof failure during the final mining phase. Meanwhile, the directional through-fracture zone formed by hydraulic fracturing is equivalent to constructing a weakening isolation zone on the path of mining stress propagation, which can block or weaken the concentrated transmission of mining stress towards the main roadway of the upper coal seam, thereby achieving active control over the deformation of the main roadway of the upper coal seam.

[0071] Furthermore, after the lower coal seam working face has been mined, if the hard roof above the goaf of the lower coal seam maintains a cantilever beam or masonry beam structure for a long period of time, it will continuously transmit high static pressure to the lower coal seam main roadway. This pressure, combined with the dynamic pressure disturbance generated by the mining of the upper coal seam, will lead to continuous convergence of the lower coal seam main roadway, floor heave, or abnormal stress on the support structure. To address this static pressure source, this invention utilizes the upper coal seam roadway as the construction space, drilling downwards through fan-shaped boreholes in the floor, allowing the borehole to reach the key layer of the hard roof above the lower coal seam. High-volume pulse fracturing is then used to form a dense fracture network. This dense fracture network weakens the integrity and bearing capacity of the key layer of the hard roof of the lower coal seam, causing the hard overhanging roof above the goaf to fully collapse. The collapsed gangue further fills the goaf with its fragmentation and expansion, forming flexible support for the overlying strata, thus eliminating or reducing the continuous high static pressure effect of the hard overhanging roof of the lower coal seam on the lower coal seam main roadway from the source.

[0072] Simultaneously, this invention optimizes the position of the upper coal seam stop line relative to the lower coal seam stop line before fracturing operations, taking into account the actual deformation characteristics of the lower coal seam main roadway. When the lower coal seam main roadway is dominated by side contraction, it indicates that horizontal stress concentration has a significant impact on the stability of the surrounding rock. By moving the upper coal seam stop line inward, the peak support pressure formed at the end of mining of the upper coal seam avoids the sensitive area of ​​the lower coal seam main roadway. When the lower coal seam main roadway is dominated by floor heave, it indicates that vertical load and floor stress transmission are dominant. By moving the upper coal seam stop line outward, the relative position between the support pressure at the end of mining of the upper coal seam and the lower coal seam main roadway is adjusted. When side contraction and floor heave are equally important, an alignment or small offset mode is adopted to achieve comprehensive coordination of horizontal stress and vertical load. Thus, the function of the stop line offset is to spatially misalign the peak support pressure with the protected main roadway, avoiding the direct superposition of high stress concentration areas on the surrounding rock of the main roadway.

[0073] By combining the above measures, upper coal seam roof fracturing mainly addresses the dynamic pressure source and its propagation path, lower coal seam roof fracturing mainly addresses the static pressure source of the hard overhanging roof in the lower coal seam goaf, stop-mining line offset mainly addresses the overlap between the peak support pressure and the spatial location of the main roadway, and monitoring and evaluation are used to dynamically verify and adjust the pressure relief effect. The four measures work together to simultaneously weaken dynamic pressure sources, eliminate static pressure sources, block stress propagation paths, and move the peak support pressure away from the protected main roadway, thus forming a comprehensive technical solution suitable for the stability control of the surrounding rock in the main roadways of the upper and lower coal seams during the final mining phase of closely spaced coal seam overlapping working faces.

[0074] Compared with using reinforced support, single roof fracturing, or simply adjusting the stop line alone, the synergistic effect of this invention is as follows: roof fracturing of the upper coal seam reduces the dynamic pressure disturbance during the final mining period; roof fracturing of the lower coal seam eliminates the continuous loading of static pressure from the hard suspended roof in the goaf; the optimized stop line avoids the peak of the support pressure coinciding with the weakened area of ​​the surrounding rock in the lower coal seam main roadway; and monitoring and evaluation ensure that the fracturing pressure relief effect can be quantitatively judged.

[0075] Example 2

[0076] This embodiment uses the mining of a coal seam group at close range in a kilometer-deep well in a coal mine as the engineering background to further describe the present invention in detail.

[0077] The mine mainly mines the upper and lower coal seams. The upper coal seam has an average thickness of 3.98m, a burial depth of -896m, and a working face length of 240m; the lower coal seam has an average thickness of 1.60m, a burial depth of -960m, and a working face length of 210m. The vertical distance between the two coal seams is approximately 64m. The lower coal seam has been mined out and sealed. The stop-mining line of the lower coal seam working face is 360m from the nearest lower coal seam main roadway, and the stop-mining line of the upper coal seam working face is 490m from the nearest upper coal seam main roadway. The horizontal distance between the stop-mining lines of the two working faces (i.e., the current rock pillar width Lp) is 108m.

[0078] Step S1: Key Layer Identification and Mechanical Testing: Core samples were drilled from the roof strata of the upper coal seam and the interlayer strata between the upper and lower coal seams for indoor physical and mechanical testing. Test results showed that a 4.0m thick layer of fine sandstone was found 10m above the roof of the upper coal seam, with a uniaxial compressive strength of 82.5MPa, a tensile strength of 6.8MPa, and an elastic modulus of 28.6GPa, identified as a key layer of the lower, hard roof of the upper coal seam. A 8.0m thick layer of medium-coarse sandstone was found 24m above the roof of the upper coal seam, with a uniaxial compressive strength of 105.3MPa, a tensile strength of 9.2MPa, and an elastic modulus of 36.4GPa, identified as a key layer of the upper, hard roof of the upper coal seam. There is a 6.0m thick sandstone layer 24m above the roof of the lower coal seam, with a uniaxial compressive strength of 78.6MPa, a tensile strength of 5.9MPa, and an elastic modulus of 24.8GPa. It is identified as the key layer of the hard roof above the lower coal seam.

[0079] Step S2, Re-determination of the stop line and optimization of its spatial relationship: Field measurements show that the deformation of the lower coal seam main roadway is mainly manifested as shrinkage of the sidewalls (convergence amount up to 620mm) and heave of the floor (heave amount 280mm). The ratio of the sidewall shrinkage to the floor heave deformation ε h ÷ε v ≈2.21, therefore α = ε is calculated. h ÷(ε h +ε v =2.21÷3.21≈0.69. Since 0.3<α<0.7, it is determined to be a composite deformation (both side compression and bottom bulging are equally important, with side compression slightly dominant).

[0080] Basic parameter values: Current interlayer rock column width L p =108m, the overburden load of the upper coal seam is q=γH≈0.025MN / m 3 ×896m≈22.4MPa, lateral pressure coefficient k=1.6, internal friction angle of interlayer rock pillar φ=32°, shear strength of surrounding rock in the lower coal seam main roadway =8.5MPa, compressive strength of the surrounding rock of the main roadway floor in the lower coal seam =32.0MPa.

[0081] First, calculate the initial fracture distance of the key layer in the hard roof of the lower coal seam. : In the formula, The initial fracture distance of this critical layer is m; The thickness of this critical layer is m; The tensile strength of this critical layer, in MPa; The load acting on this key layer is expressed in MPa. According to the key layer theory, the load on the lower key layer is the weight of the overlying rock layers controlled by it up to the upper key layer. = =0.025MN / m 3 ×14m=0.35MPa.

[0082] In the formula, The height of the rock strata between the overlying layer and the upper key layer, which is the load on the lower key layer.

[0083] =4m × (2 × 6.8MPa ÷ 0.35MPa) 1 / 2 ≈24.94m.

[0084] Based on field experience in mine pressure observation, the periodic pressure step distance is usually 1 / 2 to 1 / 3 of the initial fracture distance.

[0085] In the formula, The step distance is determined by the period of this key layer, which is the length of the fixed cantilever beam, in meters. This is an empirical coefficient, typically taken as 0.3 to 0.5. Considering the relatively hardness of the rock strata, an empirical coefficient is used. =0.5. =0.5×24.94=12.47m; For conservative calculations, the cantilever beam length is... The value is 13m.

[0086] Next, since the deformation mode of the main roadway has been determined to be a composite deformation (with equal emphasis on lateral compression and floor heave, with lateral compression slightly dominant), the composite mode is adopted to calculate the offset ΔL3.

[0087] First, calculate the optimal offset distance in both the inward and outward shift modes: Inward shift mode .

[0088] Substitute into the calculation: ΔL1=( )÷( ) ≈27.99m.

[0089] Outward mode .

[0090] Substituting the values ​​into the calculation: ΔL2 = (13m × 22.4MPa ÷ 32MPa) - 108m ≈ -98.9m.

[0091] When deformation in the main roadway of a coal seam is characterized by simultaneous contraction of the sidewalls and heave of the floor, an alignment mode or a small offset mode is adopted. The alignment / small offset mode is suitable for complex deformations and controls both horizontal and vertical stresses. Specifically, the offset ΔL3 is calculated using the following formula: In the formula, α is a weighting coefficient, which is determined based on the ratio of the shrinkage of the two sides to the deformation of the bottom drum measured on site, and the value range is 0~1. =0.69×27.99m+0.31×(-98.9m)≈-11.35m; The theoretical weighted value is -11.35m. Considering that the outward movement may affect the width of the protective coal pillar in the main roadway, the offset is rounded to 10m, that is, the upper coal seam stop mining line is moved outward by 10m.

[0092] The optimized interlayer column width is calculated using the following formula: ΔL; where, The optimized width of the interlayer rock column, in meters; the meanings of the other symbols are the same as above. ΔL = 108m + (-10m) = 98m.

[0093] Calculate the optimized mean stress of the interlayer rock column: = (22.4MPa×13m)÷98m≈2.97Mpa; =1.6×2.97MPa×tan32°≈2.97Mpa.

[0094] After optimization, the theoretical calculated stress transmitted to the lower coal seam main roadway area is ( and The average strength is approximately 2.97 MPa, which is lower than the surrounding rock strength index given in the document. =8.5MPa, =32.0MPa). This theoretically proves that optimizing the upper coal seam stop line (moving it outward by 10m) can effectively reduce the transmitted stress.

[0095] Step S3, constant flow rate water pressure fracturing of roof borehole: In the track roadway and transport roadway of the upper coal seam working face, according to the newly determined stop line position (moved 10m outward relative to the stop line of the lower coal seam), construct a fan-shaped borehole in the roof of the upper coal seam 8m behind the stop line.

[0096] Hydraulic fracturing was performed using a constant displacement high-pressure pump, with the pump injection pressure stabilized at 28~35MPa and the pump injection displacement at 80m³ / h. 3 / h, fracturing fluid is continuously injected. During the operation, the pressure curve exhibits a "rise-stabilize-sudden drop" characteristic. Water seepage from adjacent boreholes during fracturing indicates the formation of a directional, continuous fracture zone dominated by a single main fracture. Fracturing operations should be completed when the working face is at least 200m away from the predetermined stop line; since the predetermined stop line is approximately 490m from the nearest upper coal seam main roadway, the working face should be approximately 690m away from this main roadway when fracturing is completed.

[0097] Step S4, High-Volume Pulse Fracturing of Bottom Plate Drilling: Drill downwards through-layer fan-shaped bottom plate boreholes within the track roadway and transport roadway of the upper coal seam working face. Based on a vertical distance of 64m, the length of the lower coal seam working face, and a final borehole distance of 20m between adjacent boreholes, determine the detailed parameters for each borehole to ensure accurate entry of the final borehole into the 6m thick sandstone critical layer 24m above the lower coal seam. The opening position of the bottom plate boreholes is mapped to within 8m behind the lower coal seam stop line, and the line connecting the final boreholes is parallel to the working face layout direction.

[0098] Employing a high-volume pulse fracturing process, the pressure curve during fracturing exhibited periodic fluctuations with a gradually decreasing peak value, indicating the formation of a dense fracture network. This fracture zone effectively weakened the bearing capacity of the hard roof above the lower coal seam. After fracturing, the overhanging roof above the goaf of the lower coal seam fully collapsed. The measured coefficient of the collapsed gangue was 1.42, and the collapse zone height was approximately 30m, effectively filling the goaf and providing active support to the overlying strata.

[0099] Step S5, Effect Verification: Deformation monitoring points were set up in the upper and lower coal seam main roadways and monitored continuously for 6 months. In the upper coal seam main roadway, the sidewall convergence was 68mm, the floor heave was 42mm, the convergence rate was 0.4mm / d, and the floor heave rate was 0.3mm / d. Before treatment, in the lower coal seam main roadway, the sidewall convergence was 620mm, the floor heave was 280mm, the convergence rate was 2.8mm / d, and the floor heave rate was 1.5mm / d. After treatment, in the lower coal seam main roadway, the sidewall convergence was 75mm, the floor heave was 52mm, the convergence rate was 0.25mm / d, and the floor heave rate was 0.20mm / d.

[0100] Monitoring data shows that the convergence rates of both sides are less than 1 mm / d, the floor heave rate is less than 0.5 mm / d, and the cumulative convergence amount is less than 80% of the design allowable value, indicating stable anchor cable stress. Therefore, the hydraulic fracturing control of roadway deformation is deemed effective.

[0101] This embodiment verifies the effectiveness of the present invention under the condition of closely spaced coal seams stacked at a vertical distance of 64m and a mining depth of approximately 960m, and realizes the coordinated protection of the upper and lower coal seam main roadways using a single channel of the upper coal seam roadway.

[0102] The embodiments of the present invention have been described in detail above with reference to the accompanying drawings. However, the present invention is not limited thereto. Various changes that can be made within the scope of knowledge possessed by those skilled in the art without departing from the spirit of the present invention are all within the protection scope of the claims of the present invention.

Claims

1. A method for controlling fracturing of the surrounding rock in the main roadway during the final mining stage of a closely spaced coal seam superimposed working face, characterized in that, Includes the following steps: Step S1: Conduct rock mechanics tests on the roof strata of the upper coal seam and the interlayer strata between the upper and lower coal seams; The location and thickness of the key strata in the low-level hard roof and the key strata in the high-level hard roof above the upper coal seam were determined by using the key strata discrimination theory. Step S2: Based on the deformation characteristics of the lower coal seam roadway, combined with the spatial superposition relationship of the upper and lower coal seam working faces and the superposition of mining stress, determine the spatial position of the upper coal seam stop line relative to the lower coal seam stop line, establish a mechanical equilibrium model of the interlayer rock pillar, and calculate the optimal offset distance of the upper coal seam stop line relative to the lower coal seam stop line. Step S3: In the track roadway and transport roadway of the upper coal seam working face, according to the position and thickness of the key layers of the low and high hard roof determined in Step S1, and the position of the upper coal seam stop line determined in Step S2, construct fan-shaped boreholes in the roof of the upper coal seam behind the stop line to perform hydraulic fracturing on the key layers of the low and high hard roof, forming a directional through-overburden fracture zone dominated by a single main fracture. Step S4: In the track roadway and transport roadway of the upper coal seam working face, construct a fan-shaped borehole that penetrates the interlayer rock strata. The borehole ends in the key layer of the hard roof above the lower coal seam. Use high-volume pulse fracturing to hydraulically fracture the key layer of the hard roof above the lower coal seam, forming a dense fracture network. This allows the hard roof above the lower coal seam to fully collapse. The collapsing gangue is used to fill the goaf to support the roof and prevent the main roadway of the lower coal seam from being continuously affected by the high static pressure of the hard roof in the goaf. Step S5: Set up roadway deformation monitoring points in the upper and lower coal seam main roadways to monitor the roadway convergence, floor heave, and anchor cable stress changes. When the monitoring data shows that the roadway deformation rate tends to stabilize and the cumulative deformation is less than the safety set value, it is determined that the water pressure fracturing control of roadway deformation has met the standard.

2. The method for controlling fracturing of the surrounding rock in the main roadway during the final mining stage of a closely spaced coal seam superimposed working face according to claim 1, characterized in that, In step S1, the rock mechanics test specifically involves drilling rock cores for indoor physical and mechanical tests to determine the uniaxial compressive strength, tensile strength, elastic modulus, Poisson's ratio, and internal friction angle of each rock layer.

3. The method for controlling fracturing of the surrounding rock in the main roadway during the final mining stage of a closely spaced coal seam superimposed working face according to claim 1, characterized in that, In step S2, a mechanical equilibrium model of the interlayer rock column is established, including the following steps: Step 1: Determine the object of mechanical calculation: The rock strata between the upper coal seam stop mining line and the lower coal seam stop mining line, and located between the upper coal seam and the lower coal seam, are defined as interlayer rock pillars, and the horizontal distance between the upper coal seam stop mining line and the lower coal seam stop mining line is used as a spatial parameter characterizing the stress range of the interlayer rock pillars; The second step is to establish an equivalent mechanical structure: the hard roof structure that has not fully collapsed between the edge of the goaf and the stop mining line of the upper coal seam after mining is equivalent to a cantilever bearing structure that bears the load of the overlying rock strata, and the interlayer rock pillars are equivalent to a continuous rock mass structure that can transmit the mining stress of the upper coal seam. The third step is to determine the stress transmission relationship: taking the cantilever bearing structure near the coal seam stop line as the source of the upper load, we analyze the load transmission effect of its bending and sinking and end force on the upper interface of the interlayer rock column, and determine the constraint relationship between the vertical stress and the horizontal stress in the interlayer rock column. Step 4: Establish the equilibrium relationship between inter-layer rock pillars: Based on the load transmitted from the hard roof structure of the upper coal seam to the inter-layer rock pillars, the lateral constraint effect of the inter-layer rock pillars themselves, and the stress continuity at the interface between the upper and lower coal seams, determine the correspondence between the stress distribution in the inter-layer rock pillars and the relative positions of the stop mining lines of the upper and lower coal seams. Step 5: Establish the criteria for the instability of the surrounding rock in the lower coal seam roadway: Based on the measured deformation characteristics of the lower coal seam roadway, the horizontal stress corresponding to the shrinkage deformation of the two sides and the vertical stress corresponding to the heave deformation of the bottom are taken as the main control factors, and the surrounding rock strength conditions are combined to determine whether the surrounding rock in the lower coal seam roadway is in a stable state. Step 6: Determine the basis for calculating the offset of the stop mining line: Based on the inter-layer rock pillar balance relationship, stress transmission relationship and the instability criterion of the surrounding rock of the lower coal seam main roadway, determine the calculation basis for adjusting the stop mining line of the upper coal seam to move inward, outward or slightly offset relative to the stop mining line of the lower coal seam.

4. The method for controlling fracturing of the surrounding rock in the main roadway during the final mining stage of a closely spaced coal seam superimposed working face according to claim 1, characterized in that, In step S2, the calculation of the optimal offset distance between the upper coal seam stop line and the lower coal seam stop line specifically involves: Step S21: When the deformation of the main roadway of the current coal seam is mainly due to the contraction of the two sides, the inward shift mode is adopted, that is, the stop line of the upper coal seam is moved towards the goaf, and the stop line of the upper coal seam is located inside the stop line of the lower coal seam. The optimal offset distance ΔL1 is calculated by the following formula: ； In the formula, ΔL1 is the optimal distance where the upper coal seam stop mining line is located inside the lower coal seam stop mining line; L u σ is the length of the fixed-support cantilever beam in the upper coal seam; h,min L represents the shear strength of the surrounding rock in the lower coal seam main roadway. p φ is the width of the interlayer rock pillar, i.e., the horizontal distance between the stop mining lines of the upper and lower coal seams; k is the lateral pressure coefficient; q is the load of the overlying strata; φ is the internal friction angle of the interlayer rock pillar. Step S22: When the deformation of the main roadway of the lower coal seam is mainly characterized by floor heave, the outward shift mode is adopted, that is, the stop line of the upper coal seam is moved towards the main roadway, and the stop line of the upper coal seam is located outside the stop line of the lower coal seam. The optimal offset distance ΔL2 is calculated by the following formula: ； In the formula, ΔL2 is the optimal distance where the upper coal seam stop mining line is located outside the lower coal seam stop mining line; σ v,min The compressive strength of the surrounding rock in the main roadway of the lower coal seam; Step S23: When the deformation of the main roadway of the current coal seam is accompanied by both side contraction and floor heave, the alignment mode or small offset mode is adopted, and the offset amount ΔL3 is calculated according to the following formula: ； In the formula, α is a weighting coefficient, which is determined based on the ratio of the shrinkage of the two sides to the deformation of the bottom bulge measured on site, and the value range is 0~1; The offset distance ΔL is the horizontal movement distance of the upper coal seam stop line relative to its initial position. Specifically, when using the inward movement mode, ΔL = ΔL1; when using the outward movement mode, ΔL = ΔL2; when using the small offset mode, ΔL = ΔL3. When ΔL > 0, it indicates that the upper coal seam stop line moves towards the goaf; when ΔL < 0, it indicates that the upper coal seam stop line moves towards the main roadway. The width L of the inter-layer rock pillar before the adjustment of the above coal seam stop mining line p The initial width is used as the basis for optimizing and adjusting the width of the interlayer rock column based on the offset distance ΔL, resulting in the adjusted width of the interlayer rock column. It is calculated using the following formula: ； Based on the above calculation results, the upper coal seam stop line is determined to be located inside, outside or aligned with the lower coal seam stop line, ensuring that the redefined stop line is located outside or at the boundary of the support pressure influence.

5. The method for controlling fracturing of the surrounding rock in the main roadway during the final mining stage of a closely spaced coal seam superimposed working face according to claim 4, characterized in that, In step S23, the weighting coefficient α is based on the on-site measured deformation ε of the two sides. h With the deformation amount ε of the bottom drum v The ratio is determined as follows: ； When α≥0.7, it is determined that the deformation is mainly due to contraction of the two sides; When α≤0.3, it is determined that the bottom drum is dominant; When 0.3 < α < 0.7, it is determined to be a composite type.

6. The method for controlling fracturing of the surrounding rock in the main roadway during the final mining stage of a closely spaced coal seam superimposed working face according to claim 1, characterized in that, In step S3, the drilling opening position of the top plate is located within 5-10m behind the stop mining line.

7. The method for controlling fracturing of the surrounding rock in the main roadway during the final mining stage of a closely spaced coal seam superimposed working face according to claim 1, characterized in that, In step S3, the drilling and fracturing work in the upper coal seam working face track roadway and the upper coal seam working face transport roadway is completed when the working face advances to a distance of not less than 200m from the predetermined stop line.

8. The method for controlling fracturing of the surrounding rock in the main roadway during the final mining stage of a closely spaced coal seam superimposed working face according to claim 1, characterized in that, In step S3, the fan-shaped arrangement of boreholes in the upper coal seam roof has a final hole spacing of 15-20m, and the line connecting the final hole positions is parallel to the working face arrangement direction.

9. The method for controlling fracturing of the surrounding rock in the main roadway during the final mining stage of a closely spaced coal seam superimposed working face according to claim 1, characterized in that, In step S5, the safety setting is as follows: the convergence rate of the two sides of the roadway is less than 1 mm / d, the floor heave rate is less than 0.5 mm / d, the cumulative convergence amount is less than 80% of the design allowable value, and the stress fluctuation amplitude of the anchor cable is less than 30% of its yield load.

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