Deep rock burst mine drilling and anchoring synchronous pressure relief and danger elimination process method

Abstract

This invention provides a method for simultaneous stress relief and decompression in deep rockburst-prone mines, belonging to the field of coal mine rockburst prevention and control technology. By advanced detection and calibration of graded stress relief target areas, and while the integrated tunneling and anchoring machine is cutting and tunneling, it simultaneously constructs gradient stress relief holes in the sidewalls with built-in graded phase change energy dissipation bodies and biomimetic dendritic directional slotted holes in the roof formed by pulse jet molding and filled with proppant, thus constructing a three-dimensional stress relief framework with self-stabilizing progressive energy dissipation and long-term directional energy release. During construction, bottom-mounted armored temperature measurement is used. After frictional heating and geothermal gradient correction, the surrounding rock stress is inverted by coupling drilling parameters and acoustic emission signals, dynamically adjusting the stress relief parameters. After cutting, the initial support is staggered and the final support is synchronized with the static rock fracture stress relief of the floor, achieving interference-free support and stress relief coordination. This invention can achieve depth synchronization of the entire process of tunneling, stress relief, and support, completing controllable and precise stress relief without reducing tunneling efficiency.

Description

Technical Field

[0001] This invention belongs to the field of coal mine rockburst prevention and control technology, and more specifically relates to a method for simultaneous pressure relief and hazard mitigation in deep rockburst mines through tunneling and anchoring. Background Technology

[0002] As coal mining depths continue to extend, most mines have entered the 800-1500m range, and rockburst disasters have become a core bottleneck restricting the safe and efficient mining of deep coal mines. The widespread adoption of integrated tunneling and anchoring equipment has significantly increased tunneling speed, but rapid tunneling leads to a significant shortening of the stress redistribution cycle in the surrounding rock, and the accumulation rate of high elastic strain energy is far greater than in shallow mines. Under the traditional "tunneling first, then depressurization" operation mode, depressurization construction lags far behind the tunneling face, which cannot match the rapid evolution characteristics of rockburst hazards under rapid tunneling conditions, and is very likely to form an instantaneous high stress accumulation zone around the face, inducing rockburst accidents.

[0003] To address the aforementioned issues, some mines have attempted to conduct pressure relief operations in parallel with tunneling. However, existing technologies still have fundamental flaws in the coordination of tunneling, pressure relief, and support: At the pressure relief mechanism level, the energy release process of conventional equal-diameter pressure relief holes is uncontrollable; excessive coal fragmentation near the hole opening directly deteriorates the anchoring foundation, creating an inherent contradiction of "the more thorough the pressure relief, the weaker the support," and hole wall collapse in soft coal strata leads to the failure of the pressure relief channel; at the pressure relief control level, pressure relief parameters are statically designed based on prior geological data and cannot respond to dynamic stress fluctuations during tunneling. Existing stress monitoring methods rely on pre-deployed sensors, making real-time control during drilling difficult, and the accuracy of single-parameter inversion cannot meet the requirements for precise control; roof-cut pressure relief, due to the limited range of crack expansion and easy closure, has insufficient energy dissipation capacity for thick, hard roofs, while disordered cracks damage the roof's load-bearing structure, exacerbating the risk of roof subsidence.

[0004] The crux of the aforementioned deficiencies lies in the following: existing technologies treat tunneling, depressurization, and support as sequential but spatially independent processes, with depressurization merely an adjunct to tunneling rather than a collaborative system deeply integrated with tunneling and support; the function of depressurization holes is limited to "opening and releasing energy," failing to utilize the space within the holes as a controllable energy-consuming structure; stress monitoring and depressurization construction are separated into two independent stages, with the sensing results unable to provide real-time feedback to construction parameters; slotting only aims to generate cracks, neglecting the long-term maintenance of crack continuity; and depressurization and support are placed in opposition, never seeking compatible solutions from the perspectives of spatial avoidance and temporal coordination. This "tunneling-depressurization-support" three-part separation technical architecture is the root cause of a series of problems such as delayed depressurization, uncontrolled energy release, and support-depressurization conflicts.

[0005] Therefore, developing a new decompression and hazard mitigation technology that can achieve coordinated excavation, unloading, and support depths, controllable and adjustable decompression processes, and compatible support and unloading performance is a technical challenge that urgently needs to be solved in the field of rapid excavation and anti-rockburst prevention in deep rockburst mines. Summary of the Invention

[0006] To address the above problems, the present invention aims to provide a method for simultaneous stress relief and decompression in deep mines prone to rockburst. This method utilizes a technical system of stepped positioning and graded phase change energy dissipation, long-term stress relief through pulsed directional dendritic cutting, multi-source corrected stress inversion, and time-sequential staggered support coordination. This system achieves deep, simultaneous parallel operation of tunneling, stress relief, and support processes, enabling controllable and precise stress relief without reducing tunneling efficiency. Simultaneously, it solves common industry problems such as instability of the stress relief borehole wall, insufficient accuracy of stress inversion, easy closure and failure of cuttings, and mutual interference between support and decompression, thus comprehensively improving the effectiveness of rockburst prevention and control in deep mines.

[0007] To achieve the above objectives, the present invention employs the following technical solution: This application provides a method for simultaneous pressure relief and hazard mitigation through tunneling and anchoring in deep rockburst mines, comprising the following steps: S1: Before tunneling, detect the geological parameters and stress distribution of the surrounding rock ahead, classify the impact hazard level, mark the graded stress relief target areas corresponding to the roof, coal seam and floor respectively, and set stress control intervals including the first threshold and the second threshold for each target area. S2: A single-cycle operation is performed by a tunneling and anchoring machine. While the cutting section is cutting the roadway, a machine-mounted side drilling rig is used to construct gradient energy-dissipating and pressure-relieving holes on both sides of the roadway and implant phase change energy-dissipating bodies. A machine-mounted top-advanced drilling rig is used to construct biomimetic dendritic directional cutting holes in the roof in front of the face. The gradient energy-dissipating and pressure-relieving holes are axially variable diameter structures, and phase change energy-dissipating bodies with yield thresholds gradually increasing from shallow to deep are set in sections inside to form graded progressive energy-dissipating channels. The biomimetic dendritic directional cutting holes are formed by water jet cutting to create main and branch seams and injecting supporting particles to construct a directional fracture pressure-relieving network. S3: During the pressure relief drilling process, drilling response parameters and surrounding rock response signals are collected. After interference correction, the stress concentration coefficient of the surrounding rock in the current construction area is obtained by the stress concentration coefficient inversion method. The inverted stress concentration coefficient is compared with the stress control interval. When the stress concentration coefficient is higher than the first threshold or lower than the second threshold, the construction parameters of the subsequent pressure relief holes and slotted holes that have not yet been constructed in this cycle are adjusted accordingly. S4: After the single-cycle cutting and pressure relief of the side and top plate are completed, temporary initial support is constructed first to maintain the stability of the surrounding rock, and then permanent support anchoring and static directional rock-cracking pressure relief of the bottom plate are carried out simultaneously; the anchoring section of the permanent anchor rod avoids the borehole expansion section of the gradient energy dissipation pressure relief hole. S5: After completing a single cycle of pressure relief and support, the tunneling and anchoring machine advances one cycle of progress. The pressure relief effect is checked at a set distance behind the face, and the check result is used as the basis for correction of the target area calibration of the next cycle of graded pressure relief. After the check is qualified, the next working cycle begins.

[0008] In one optional embodiment, the gradient energy dissipation and pressure relief hole is provided axially with an enlarged orifice section, a narrowed middle section, and an original orifice section in sequence; the diameter of the enlarged orifice section is 1.5 to 2 times the diameter of the original orifice section, and the length is 1.5 to 3 m; the diameter of the narrowed middle section is 0.6 to 0.8 times the diameter of the original orifice section, and the length is 1 to 2 m; an annular support step is formed at the junction of the enlarged orifice section and the narrowed middle section, and at the junction of the narrowed middle section and the original orifice section.

[0009] In one optional embodiment, an embedded limiting ring is provided at the orifice of the gradient energy dissipation and pressure relief hole, and the end of the phase change energy dissipator has an annular flange that abuts against the annular support step and the retaining ring to achieve axial fixation; the end face of the energy dissipator is prefabricated with an annular deformation induction groove, and the yield threshold is gradually increased from shallow to deep, so that the energy dissipates in sequence according to the hole depth when under pressure.

[0010] In one optional embodiment, the advance distance of the biomimetic dendritic directional slotting borehole is not less than 5m, and the elevation angle is 10°~20°. During the drilling process, the drill rod has an inclination sensor built in it to correct the drilling trajectory in real time. After drilling to the design depth, a pulsed high-pressure water jet is used to cut along the key layer of the top plate to form the main slot. The length of the main slot in a single hole is not less than 10m. Then, multiple branch slots are cut along the axis of the main slot at an angle of 25°~35°. After the slotting is completed, high-pressure air is used as a transport carrier to carry the supporting particles into the fissure. The supporting particles are high-strength ceramic particles that support the fissure surface to resist the closure of the confining pressure.

[0011] In an optional implementation, the method for inverting the stress concentration factor includes: A baseline database corresponding to drilling pressure, rotation speed, flushing fluid flow rate, and frictional heat generation is established in advance through calibration holes on the working face. Drilling parameters are collected in real time and the frictional heat generation component is deducted. Then, the geothermal gradient is corrected according to the borehole depth to obtain the stress-induced temperature increment. Drilling torque, corrected temperature increment, and acoustic emission event frequency are coupled according to weights to establish a mapping model to calculate the stress concentration factor of the surrounding rock at the current location. When the stress concentration factor is higher than the first threshold, the spacing of the pressure relief holes is reduced, the length of the enlarged section is increased, and the energy consumption level of the energy-consuming body is improved, while the cutting branch density is increased. When the stress concentration factor is lower than the second threshold, the spacing of the pressure relief holes is increased and the length of the enlarged section is shortened, while the cutting branch density is reduced. The first threshold is greater than the second threshold.

[0012] In an optional implementation, the acquisition of drilling response parameters and surrounding rock response signals includes: The drilling torque and propulsion resistance are collected in real time by the armored sensor group built into the front end of the drill pipe, the temperature of the drill cuttings at the bottom of the hole is directly collected by the armored platinum resistance temperature sensor, and the micro-fracture signal of the surrounding rock is collected simultaneously by the miniature acoustic emission sensor in the roadway side; at the same time, the drilling pressure, rotation speed and flushing fluid flow rate are also collected. The interference correction includes frictional heat generation correction and geothermal gradient correction; The frictional heat generation correction is as follows: a reference library corresponding to drilling pressure, rotation speed, flushing fluid flow rate and drill bit frictional heat generation is established in advance through the calibration hole of the working face; during construction, drilling pressure, rotation speed and flushing fluid flow rate are collected in real time and the corresponding frictional heat generation is deducted. The geothermal gradient correction is performed by subtracting the original geothermal temperature increment with depth based on the borehole depth and the regional geothermal gradient value.

[0013] In one optional implementation, the temporary initial support adopts quick-installation end anchor bolts, with a single bolt installation time not exceeding 90 seconds. The pressure relief operation of the base plate is started immediately after the initial support is completed. The permanent support anchoring and the static directional rock-cracking pressure relief of the base plate are carried out in parallel. The pressure relief short holes are constructed on the base plate using hydraulic static cracking technology. The rock-cracking depth of the pressure relief short holes is 3~6m, and the hole spacing is 1.5~2.5m. The resulting loosening zone of the base plate is spatially connected with the gradient energy-dissipating pressure relief holes on both sides.

[0014] In an optional implementation, step S2 includes: The parallel overlap of the three actions performed by the cutting department—cutting the roadway, constructing gradient energy-consuming and pressure-relieving holes on both sides of the roadway using a machine-mounted side drilling rig, and constructing biomimetic dendritic directional cutting holes in the roof of the face using a machine-mounted top advanced drilling rig—is no less than 80%, and the total duration of these three actions does not exceed 1.15 times the cutting operation duration.

[0015] In one optional implementation, the pressure relief effect is verified by a combination of drill cuttings method, online stress monitoring and borehole wall imaging. When the amount of drill cuttings is below the critical value, the stress monitoring data shows a downward trend and the development of borehole wall cracks is within the design range, the pressure relief is deemed qualified.

[0016] In one optional implementation, the advance of one cycle is 1.5~2.5m, the pressure relief and verification position is located 8~15m behind the face, and the biomimetic dendritic directional cutting borehole always maintains a lead distance of not less than 3m.

[0017] As can be seen from the above technical solutions, the present invention has the following advantages: This invention constructs a dual self-stabilizing progressive energy dissipation mechanism combining stepped positioning and graded phase transition. By utilizing the borehole diameter-changing steps and limiting retaining rings, the axial direction of the energy dissipation body is completely fixed, structurally solving the problem of energy dissipation body failure due to axial movement under high stress. Combined with deformation-inducing grooves and progressively increasing yield thresholds, energy is released in an orderly manner from shallow to deep, completely avoiding skip-stage failure. The energy dissipation body also has a wall protection function, solving the industry pain point of easy collapse of pressure relief holes in soft coal formations, and significantly improving the long-term stability of pressure relief effect.

[0018] The pulse-directed dendritic cutting and proppant filling long-term pressure relief technology proposed in this invention ensures the cutting accuracy through drilling trajectory correction and avoids the loss of control of the fracture network caused by original fractures. The ceramsite proppant filling fundamentally solves the problem of rapid cutting closure under deep high ground stress and significantly extends the effective working period of the cutting. The dendritic structure not only expands the pressure relief range, but also achieves controllable fracture direction, avoiding the damage of disordered fractures to the roof load-bearing structure, and achieving a balance between roof pressure relief and structural stability.

[0019] This invention establishes a multi-source stress inversion system with bottom hole temperature measurement and dual correction. It directly collects the temperature of fresh drill cuttings at the bottom of the hole to avoid heat dissipation errors. It removes interfering factors through dual correction of frictional heat generation and geothermal gradient. Combined with a weighted coupling model, it significantly improves the accuracy of stress inversion while drilling and solves the problems of large interference and high inversion error of traditional temperature parameters, providing reliable data support for dynamic control.

[0020] This invention employs a staggered initial support and synchronized final support timing strategy, combined with a hydraulic static cracking process for the base plate, to completely resolve the performance conflict between pressure relief and support operations. This achieves a high degree of parallelism in process timing while avoiding the negative impact of vibration and water immersion on anchoring quality. At the same time, the spatially staggered design between the anchoring section and the pressure relief zone ensures structurally that the support bearing capacity is not degraded by the pressure relief operation, achieving a deep synergy between pressure relief and surrounding rock stability.

[0021] This invention achieves deep, synchronous, and parallel operation of the entire process of tunneling, depressurization, and support. The single-cycle operation time is basically equivalent to that of simple tunneling and anchoring operations, which completely solves the problem of traditional depressurization occupying tunneling time. At the same time, it constructs a three-dimensional depressurization system with long-term pre-splitting of the roof, gradient energy consumption of the coal seam, and static rock fracturing of the floor, which can simultaneously resolve multiple impact risks and perfectly meet the anti-impact requirements of safe and efficient tunneling in deep mines. Attached Figure Description

[0022] To more clearly illustrate the technical solution of the present invention, the accompanying drawings used in the description will be briefly introduced below. Obviously, the accompanying drawings described below are only some embodiments of the present invention. For those skilled in the art, other drawings can be obtained based on these drawings without creative effort.

[0023] Figure 1 A schematic diagram of the process for simultaneous decompression and hazard mitigation in deep rockburst mines provided in this application.

[0024] Figure 2 A schematic diagram of another deep rockburst mine tunneling and anchoring synchronous pressure relief and hazard mitigation process provided in this application.

[0025] Figure 3 A flowchart illustrating the multi-source coupled stress inversion and dynamic control method provided in this application. Detailed Implementation

[0026] The various embodiments of this disclosure will be described more fully in the following detailed description of the specific steps of the synchronous decompression and hazard mitigation process for deep rockburst mines. This disclosure may have various embodiments, and adjustments and changes may be made therein. However, it should be understood that there is no intention to limit the various embodiments of this disclosure to the specific embodiments disclosed herein, but rather this disclosure should be understood to cover all adjustments, equivalents, and / or alternatives falling within the spirit and scope of the various embodiments of this disclosure.

[0027] In the following, the terms “comprising” or “may include”, which may be used in various embodiments of this disclosure, indicate the presence of the disclosed functions, operations, or elements, and do not limit the addition of one or more functions, operations, or elements. Furthermore, as used in various embodiments of this disclosure, the terms “comprising,” “having,” and their cognates are intended only to indicate a particular feature, number, step, operation, element, component, or combination of the foregoing, and should not be construed as primarily excluding the presence of one or more other features, numbers, steps, operations, elements, components, or combinations of the foregoing, or the possibility of adding one or more combinations of the foregoing.

[0028] The technical solutions of the embodiments of the present invention will be clearly and completely described below with reference to the accompanying drawings. Obviously, the described embodiments are only some embodiments of the present invention, and not all embodiments. Based on the embodiments of the present invention, all other embodiments obtained by those skilled in the art without creative effort are within the scope of protection of the present invention.

[0029] Please see Figure 1 The diagram shows a flowchart of a method for simultaneous stress relief and hazard mitigation in deep rockburst mines, as described in a specific embodiment. The method includes: S1: Before tunneling, detect the geological parameters and stress distribution of the surrounding rock ahead, classify the impact hazard level, mark the graded stress relief target areas corresponding to the roof, coal seam and floor, and set stress control intervals including the first threshold and the second threshold for each target area.

[0030] In specific implementations, advance detection is conducted before tunneling operations. Geological parameters and stress distribution of the surrounding rock ahead are obtained using equipment such as airborne ground-penetrating radar or advanced borehole stress gauges. In actual use, two detection methods can be used together for mutual verification. Based on the detection results, the area ahead is classified into rockburst hazard levels according to the rockburst hazard classification standard. Then, combined with stress distribution characteristics, the spatial range of graded stress relief target areas is marked in the roof, coal seam sides, and floor. Based on this, the target areas are divided into strong stress relief zones, conventional stress relief zones, and weak stress relief zones according to stress levels, and a stress control interval including a first threshold and a second threshold is pre-set for each target area. Before the formal construction of each working face, 2-3 calibration boreholes are drilled to complete the calibration of localized parameters, providing a benchmark for subsequent stress inversion.

[0031] S2: A single-cycle operation is performed using a tunneling and anchoring integrated machine. While the cutting section is cutting the roadway, a machine-mounted sidewall drilling rig is used to construct gradient energy-dissipating and pressure-relieving holes on both sides of the roadway and implant phase change energy-dissipating bodies. A machine-mounted top-advanced drilling rig is used to construct biomimetic dendritic directional cutting holes in the roof in front of the face. The gradient energy-dissipating and pressure-relieving holes are axially variable diameter structures, and phase change energy-dissipating bodies with yield thresholds gradually increasing from shallow to deep are set in sections inside to form graded progressive energy-dissipating channels. The biomimetic dendritic directional cutting holes are formed by water jet cutting to create main and branch seams and injecting supporting particles to construct a directional fracture pressure-relieving network.

[0032] In a specific implementation, after the integrated tunneling and anchoring machine enters single-cycle operation, while the cutting section performs tunnel cutting, the onboard sidewall drilling rig constructs gradient energy-dissipating and pressure-relief holes on both sides of the tunnel. These pressure-relief holes have a variable diameter structure along the axial direction, consisting of a borehole enlargement section, a middle necking section, and the original borehole section. The diameter of the borehole enlargement section is 1.5 to 2 times the diameter of the original borehole section, and its length is 1.5 to 3 meters; the diameter of the middle necking section is 0.6 to 0.8 times the diameter of the original borehole section, and its length is 1 to 2 meters. An annular support step is formed at the junctions of the enlargement section and the necking section, and at the junctions of the necking section and the original borehole section. In actual construction, the above three-section structure is usually formed by first drilling the original hole and then enlarging it in sections. After construction is completed, the phase change energy-dissipating body is sequentially fed into each section of the hole using a pusher rod. The annular flange at the end of the energy-dissipating body abuts against the annular support step, and an embedded limit ring is installed at the borehole opening to achieve full axial fixation. Each section of the energy-consuming body has a pre-fabricated annular deformation induction groove on its end face. The yield threshold is gradually increased from shallow to deep. When under pressure, the energy is consumed in sequence according to the hole depth, forming a graded and progressive energy-consuming channel.

[0033] Simultaneously with the drilling of the ribs, a biomimetic dendritic directional cutting borehole is constructed into the top slab in front of the face using an onboard top-mounted advanced drilling rig. The advance distance of the borehole is no less than 5m, and the elevation angle is 10°~20°. During the drilling process, the drill rod's built-in tilt sensor corrects the drilling trajectory in real time. After the borehole reaches the designed depth, a pulsed high-pressure water jet is used to first cut along the key layer of the top slab to form the main seam, with a single main seam length of no less than 10m. Then, multiple levels of branch seams are cut along the axial direction of the main seam at intervals of 25°~35°. After the seams are cut, high-pressure air is used as a transport carrier to inject high-strength ceramic proppant into the fractures to support the fracture surface against confining pressure closure. The above three actions—cutting, rib drilling, and top slab cutting—are carried out in parallel, with an overlap of no less than 80%, and the total time of the three actions does not exceed 1.15 times the cutting operation time.

[0034] S3: During the pressure relief drilling process, drilling response parameters and surrounding rock response signals are collected. After interference correction, the stress concentration coefficient of the surrounding rock in the current construction area is obtained by the stress concentration coefficient inversion method. The inverted stress concentration coefficient is compared with the stress control interval. When the stress concentration coefficient is higher than the first threshold or lower than the second threshold, the construction parameters of the subsequent pressure relief holes and slotted holes that have not yet been constructed in this cycle are adjusted accordingly.

[0035] In a specific implementation, during the pressure relief drilling process, drilling torque and propulsion resistance are collected in real time by an armored sensor group built into the drill pipe front end, and the temperature of the drill cuttings at the bottom of the hole is directly collected by an armored platinum resistance temperature sensor. Simultaneously, micro-fracture signals of the surrounding rock are collected by a miniature acoustic emission sensor on the roadway side. Drilling pressure, rotation speed, and flushing fluid flow rate are also collected for subsequent frictional heat generation correction. The collected signals are sent to the data processing unit, where frictional heat generation correction and geothermal gradient correction are performed first. Frictional heat generation correction is based on a pre-established benchmark library of corresponding drilling pressure, rotation speed, flushing fluid flow rate, and frictional heat generation established through calibration holes on the working face. During construction, drilling pressure, rotation speed, and flushing fluid flow rate are collected in real time, and the corresponding frictional heat generation is deducted. Geothermal gradient correction deducts the original geothermal temperature increment with depth based on the borehole depth and the regional geothermal gradient value. After completing the dual correction, the drilling torque, corrected temperature increment, and acoustic emission event frequency are coupled according to weights and substituted into the calibrated mapping model to obtain the stress concentration factor of the surrounding rock in the current area.

[0036] The obtained stress concentration factor is compared in real time with the stress control range set for the target area in S1. If it is higher than the first threshold, the spacing between subsequent pressure relief holes is reduced, the length of the enlarged hole section is increased, and the energy dissipation level of the energy-consuming body is improved, while the density of slit branches is increased; if it is lower than the second threshold, the spacing between pressure relief holes is increased, the length of the enlarged hole section is shortened, and the energy dissipation level is reduced, while the density of slit branches is decreased. The adjustment is limited to subsequent pressure relief holes and slit holes that have not yet been constructed in this cycle.

[0037] S4: After the single-cycle cutting and pressure relief of the side and top plate are completed, temporary initial support is constructed first to maintain the stability of the surrounding rock, and then permanent support anchoring and static directional rock-cracking pressure relief of the bottom plate are carried out simultaneously; the anchoring section of the permanent anchor rod avoids the borehole expansion section of the gradient energy-dissipating pressure relief hole.

[0038] In a specific implementation, after the single-cycle cutting and decompression of the sidewalls and roof are completed, temporary initial support construction is carried out first. The temporary initial support uses quick-installing end anchor bolts, with a single bolt installation time not exceeding 90 seconds, quickly forming temporary constraints on the surrounding rock to prevent sidewall spalling or roof collapse in subsequent operations. After the temporary initial support is completed, permanent support anchoring and static directional rock-breaking decompression of the floor slab are simultaneously initiated.

[0039] During permanent support construction, the top and side anchors are arranged according to design parameters. The anchoring section of the permanent side anchors avoids the enlarged section of the gradient energy dissipation and pressure relief holes, and the entire anchoring section is located within the intact coal body surrounding the original hole section below the necking section to ensure stable and reliable anchoring force. Static directional rock fracturing and pressure relief of the floor is achieved using hydraulic static fracturing technology. Short pressure relief holes are constructed on the floor, with a fracturing depth of 3-6m and a hole spacing of 1.5-2.5m. High-pressure expansion fracturing forms a loosening zone in the floor, which spatially connects with the gradient energy dissipation and pressure relief holes on both sides, forming a continuous pressure relief area surrounding the roadway.

[0040] S5: After completing a single cycle of pressure relief and support, the tunneling and anchoring machine advances one cycle of progress. The pressure relief effect is checked at a set distance behind the face, and the check result is used as the basis for correction of the target area calibration of the next cycle of graded pressure relief. After the check is qualified, the next working cycle begins.

[0041] In a specific implementation, after completing a single cycle of pressure relief and support, the tunneling and anchoring machine advances one cycle's advance, which is 1.5~2.5m. At a distance of 8~15m behind the face, a comprehensive verification of the pressure relief effect is performed using a combination of drill cuttings method, online stress monitoring, and borehole wall imaging. When the amount of drill cuttings is below a critical value, the stress monitoring data shows a downward trend, and the borehole wall crack development is within the design range, the pressure relief is deemed qualified. The verification data is used as the correction basis for the graded pressure relief target area calibration in the next cycle S1, updating the target area division or stress control interval setting, and then proceeding to the next work cycle. If the verification fails, additional pressure relief holes are constructed until the standard is met before proceeding to the next cycle. Throughout the process, the biomimetic dendritic directional slotted drilling maintains a lead distance of no less than 3m to ensure that the top plate in front of the face is always in a pre-pressure relief state.

[0042] In this embodiment, addressing the fundamental contradiction of pressure relief lagging behind tunneling under rapid excavation conditions in deep rockburst mines, a holistic technical approach of time-space-depth coordination of excavation, pressure relief, and support is proposed: Temporally, pressure relief drilling and cutting operations are synchronized and parallel, eliminating the pressure relief lag window; Energy-wise, the disordered instantaneous release of energy in traditional constant-diameter boreholes is transformed into sequential energy consumption by graded yield bodies within variable-diameter boreholes, making the energy release process controllable; Spatially, the anchoring section avoids the fractured zone at the borehole opening, preventing pressure relief from deteriorating the support foundation; Directional main-branch cuts are used to guide the orderly expansion of fractures in the roof and inject support bodies to maintain long-term conductivity, preventing disordered fractures from damaging the load-bearing structure. Based on this, data is collected while drilling using sensors built into the drill pipe, and stress is inverted in real time. Subsequent drilling parameters are dynamically adjusted accordingly, enabling the pressure relief intensity to adaptively match stress fluctuations, ultimately forming a dynamic operation mode that adapts to both pressure relief construction and the surrounding rock condition.

[0043] Furthermore, as a refinement and extension of the specific implementation of the above embodiments, in order to fully illustrate the specific implementation process in this embodiment, another method for simultaneous decompression and hazard relief in deep rockburst mines is provided.

[0044] like Figure 2 As shown, the method includes the following steps: Step 1: Advanced Detection and Target Area Determination: Before tunneling operations, airborne ground-penetrating radar combined with advanced borehole stress gauges is used to conduct detailed detection of the surrounding rock ahead, obtaining geological parameters such as coal seam thickness, roof lithology and stratigraphy, and the magnitude and direction of ground stress. The rockburst hazard level of the area ahead is classified according to the rockburst hazard classification standard. Based on the stress distribution characteristics, the spatial range of graded stress relief target areas corresponding to the key roof strata, coal seam sidewalls, and high-stress areas of the floor are determined. The target areas are divided into strong stress relief zones, conventional stress relief zones, and weak stress relief zones according to stress levels, and the stress relief intensity threshold and parameter benchmarks for each zone are clearly defined.

[0045] Step 2, Synchronous and Parallel Excavation and Unloading Operation: The tunneling and anchoring machine enters a single-cycle operation process. While the cutting section performs tunnel cutting operations according to the set cross-section, the side drilling rigs on the left and right sides of the machine simultaneously construct gradient energy-consuming pressure relief holes in the corresponding pressure relief target areas on both sides of the tunnel. After the drilling is completed, the phase change energy-consuming body is immediately implanted. The advanced drilling rig on the top of the machine simultaneously constructs biomimetic branch-shaped directional cutting holes in the roof target area in front of the face.

[0046] The gradient energy dissipation and pressure relief hole is constructed along the borehole axis with three variable-diameter sections: an enlarged section at the orifice, a narrowed section in the middle, and the original section at the bottom. The junctions between the enlarged and narrowed sections, and between the narrowed and original sections, form natural annular support steps. Each section is equipped with an independent columnar phase-change energy dissipator. The end of the energy dissipator has an annular flange; after implantation, the flange abuts against the corresponding step surface. Combined with a limiting retaining ring embedded at the orifice, this achieves full axial fixation of the energy dissipator, completely preventing movement, misalignment, or extrusion of the energy dissipator under high stress. The yield threshold of the energy dissipator in each section gradually increases from shallow to deep, with the lowest yield threshold at the orifice, followed by the narrowed section, and the highest at the bottom. An annular deformation induction groove is pre-fabricated on the pressure-bearing end face of each energy dissipator section to guide plastic deformation in a preset sequence. High-strength rigid spacers are installed between sections to ensure uniform stress transmission while preventing inter-section embedding and damage. When under pressure, the energy-consuming body at the orifice section yields first and dissipates energy. As the stress continues to rise, the energy-consuming bodies in the middle and deep parts are activated in sequence, forming a progressive energy-consuming channel with graded yielding, ensuring the orderly and controllable release of energy.

[0047] The phase change energy dissipator employs a composite structure with a paraffin phase change core material supported by a cement-based porous carrier, encased in a flame-retardant and antistatic engineering plastic shell. This structure meets mining safety standards while preventing leakage of the phase change material and erosion from coal and water, maintaining long-term performance stability. The energy dissipator also acts as a borehole wall support, effectively preventing borehole collapse under high stress and maintaining the long-term stability of the pressure relief channel. Furthermore, the material degrades slowly in the mining environment, without affecting long-term stress release.

[0048] The biomimetic dendritic directional slotting drilling employs pulsed high-pressure water jet slotting technology. The drill rod incorporates tilt and trajectory sensors to correct the drilling direction and slotting angle in real time. When encountering primary bedding or joints, it automatically adjusts the jet pressure and feed rate to ensure the slotting extends along the designed path, preventing disordered fracture propagation. The cutting creates a dendritic structure with a main slot and multiple branch slots: the main slot extends along the bedding plane of the hard, critical layer of the roof, cutting off the stress transmission path of the rock strata; the branch slots are at a predetermined angle to the main slot and are evenly distributed along the axial direction of the main slot, directionally expanding the fracture network, significantly increasing the energy release range, and simultaneously guiding the fractures to develop along a preset direction, preventing disordered fractures from damaging the overall load-bearing structure of the roof. As an example, the advance distance of the biomimetic dendritic directional slotted drilling is not less than 5m, and the elevation angle is 10°~20°; the drill rod has a built-in tilt sensor to correct the trajectory in real time, the main slot formed by pulse jet cutting extends along the key layer of the top plate, and the length of the main slot in a single hole is not less than 10m; the branch slots are at an angle of 25°~35° to the main slot and are distributed at intervals along the axial direction of the main slot.

[0049] After the cut is formed, high-strength ceramic proppant is evenly delivered into the cut using high-pressure air via a drill pipe, filling the space between the upper and lower fracture surfaces to form a continuous point-support system. This effectively resists fracture closure caused by high ground stress and maintains the long-term conductivity of the cut's pressure relief channel. At the junction of the main fracture and branch fracture, a dilatation cutting technique is used to create a diameter-enlarging node, dispersing stress concentration at the fracture opening and preventing premature closure failure at the junction. The cut holes in the top slab always maintain a safe distance ahead of the facing direction, achieving pre-fracture pressure relief.

[0050] Specifically, the outer shell of the phase change energy dissipator is made of mining flame-retardant and anti-static PVC material, and the yield threshold of the internal composite phase change material can be precisely controlled by the material ratio and the depth of the induction groove. After the drilling is completed, the energy dissipator and the partition are pushed into the corresponding positions in the hole by the push rod of the drilling rig, and the limiting retaining ring is driven into the hole to complete the fixation.

[0051] Specifically, the particle size of the ceramic proppant is 0.5~1mm, the injection volume per hole is determined according to the cut volume, and the injection pressure is lower than the cut pressure to avoid proppant entering the original cracks and causing waste.

[0052] Step 3: Multi-source Coupled Stress Sensing and Parameter Control. This step specifically discloses a method for multi-source coupled stress inversion and dynamic control. For example... Figure 3 As shown, the specific process is as follows: During pressure relief drilling, a built-in armored sensor array at the drill pipe tip collects drilling load parameters such as drilling torque and propulsion resistance in real time. Simultaneously, an armored platinum resistance temperature sensor directly collects the temperature of fresh drill cuttings at the bottom of the hole, avoiding temperature distortion caused by heat dissipation from the borehole opening. At the same time, a miniature acoustic emission sensor in the roadway side collects micro-fracture signals in the surrounding rock.

[0053] Before stress inversion, a dual interference correction is performed: First, frictional heat generation correction. Tests are conducted in advance through calibration holes on the working face to establish a baseline library of drill bit frictional heat generation under different drilling pressures, rotation speeds, and flushing fluid flow rates. Drilling parameters are collected in real time during construction, and the frictional heat generation under the corresponding working conditions is deducted. Second, geothermal gradient correction. Based on the borehole depth and the regional geothermal gradient value, the increase in the original geothermal temperature with depth is deducted, and finally, the temperature increase caused only by stress concentration is obtained.

[0054] The corrected temperature increment is coupled with drilling torque and acoustic emission event frequency according to weights, with drilling torque as the main weight parameter and temperature increment and acoustic emission frequency as correction parameters. A localized mapping model is established in advance through on-site calibration, and the stress concentration factor of the surrounding rock in the current construction area is obtained by substituting the real-time parameters into the model.

[0055] Based on the threshold range of the stress concentration coefficient, subsequent pressure relief construction parameters are dynamically adjusted: when the stress concentration coefficient is higher than the set first threshold, it is identified as a high-stress zone, and the spacing between pressure relief holes is reduced, the length of the enlarged hole section is increased, and a phase changer with a higher energy dissipation level is used. At the same time, the branch density of the top plate cuts is increased to enhance the pressure relief intensity. When the stress concentration coefficient is lower than the set second threshold, it is identified as a low-stress zone, and the spacing between pressure relief holes is increased, the length of the enlarged hole section is shortened, and the energy dissipation level is reduced. At the same time, the branch density of the cuts is reduced to reduce the amount of construction work while ensuring the pressure relief effect. Through closed-loop control of drilling, testing, and adjustment, dynamic and precise matching between pressure relief parameters and the stress state of the surrounding rock is achieved.

[0056] Specifically, in step 3, the first threshold and the second threshold are preset according to the mine's impact hazard level. The threshold for mines with strong impact hazard is lower than that for mines with medium impact hazard. Before the formal construction of each working face, 2 to 3 calibration holes are constructed to complete the localization parameter calibration.

[0057] Step 4, Staggered timing of support and unloading and pressure relief of the bottom plate: After the single-cycle cutting operation and the pressure relief construction of the side and top plate are completed, the timing strategy of "staggered initial support and simultaneous final support" is adopted: first, the temporary initial support anchor bolts are quickly constructed to form a temporary constraint on the surrounding rock, maintain the initial stability of the surrounding rock after cutting, and avoid the occurrence of side and top collapse in subsequent operations; after the initial support is completed, the permanent support anchoring operation and the static pressure relief operation of the bottom plate are started simultaneously.

[0058] The floor slab utilizes a hydraulic static fracturing process to construct short pressure relief holes. High-pressure hydraulic fracturing of the rock mass creates a directional loosening zone, without any blasting vibrations or harmful gases, thus avoiding disturbance to the support anchoring. The anchoring depth of the permanent sidewall anchors matches the structure of the gradient energy-dissipating pressure relief holes, ensuring that the anchoring section is entirely located within the intact coal seam surrounding the original hole section below the necking section. This completely avoids the fractured area and phase change energy-dissipating body range of the hole enlargement section, guaranteeing stable and reliable anchoring force and achieving compatibility between pressure relief and support performance. The floor slab loosening pressure relief zone and the gradient energy-dissipating pressure relief holes on both sides are spatially interconnected, forming a continuous pressure relief zone surrounding the roadway, uniformly transferring high stress from the roadway perimeter to the deeper layers of the surrounding rock.

[0059] For example, in this step, the temporary initial support uses quick-installation end anchor bolts, with a single bolt installation time not exceeding 90 seconds. The floor rock-cracking operation begins immediately after the initial support is completed. The permanent support anchoring and floor static rock-cracking are carried out in parallel. The floor uses hydraulic static cracking technology to construct short pressure relief holes, with a cracking depth of 3-6m and a hole spacing of 1.5-2.5m. The resulting loosened zone in the floor connects spatially with the pressure relief holes on both sides. The anchoring section length of the permanent anchor bolts on the sides is not less than 1m, and the entire anchoring section is located within the intact coal seam area below the necking section of the gradient energy-dissipating pressure relief hole. Phase change energy-dissipating bodies do not intrude into the anchoring section area.

[0060] In addition, it should be noted that in steps 2 and 4, the overlap between the cutting operation, the side pressure relief drilling construction, and the top plate cutting construction shall not be less than 80%, and the temporary initial support, permanent support, and bottom plate pressure relief shall be connected and carried out in parallel. The total time of all procedures in a single cycle shall not exceed 1.15 times the cutting operation time.

[0061] Step 5, Lag Verification and Cyclic Advancement: After completing a single cycle of pressure relief and support operations, the tunneling and anchoring machine advances one cycle's advance. At a predetermined distance behind the face, a comprehensive verification of the pressure relief effect is performed using a combination of drill cuttings method, online stress monitoring, and borehole wall imaging. This verifies the degree of coal stress release, the development range of the plastic zone, and borehole wall stability. If the verification results meet the pressure relief index requirements, the pressure relief for this cycle is deemed qualified, and the next operation cycle continues. If the verification fails, additional pressure relief holes are constructed until the requirements are met.

[0062] Specifically, the single-cycle advance is set to 1.5~2.5m, the pressure relief check position lags behind the face by 8~15m, and the top plate branch cut hole always maintains a lead distance of not less than 3m.

[0063] The present invention will be further described below with reference to specific embodiments. These embodiments are only used to illustrate the implementation effects of the present invention and do not constitute a limitation on the scope of protection of the present invention. Embodiment 1 corresponds to the conventional application scenario of mines with medium impact hazard, Embodiment 2 corresponds to the enhanced application scenario of mines with thick and hard roofs and strong impact hazard, and Embodiment 3 corresponds to the special application scenario of mines near fault structures.

[0064] Example 1: This embodiment involves an industrial test conducted in a deep mine with moderate impact hazard. The mine has a mining depth of 860m, ​​an average main coal seam thickness of 3.2m, a uniaxial compressive strength of 12.5MPa, an immediate roof of 4.5m thick sandy mudstone, a secondary roof of 8m thick fine sandstone, a maximum horizontal principal stress of 22.3MPa, an initial stress concentration factor of 1.42, and is classified as having moderate impact hazard. The mine previously employed a tunneling and anchoring machine combined with delayed large-diameter borehole pressure relief. A single cycle operation time was 45 minutes, the average monthly advance was 178m, the average approach distance between the roadway sides was 210mm, the anchor bolt anchoring strength qualification rate was approximately 89%, and an average of 7.2 impact hazard warnings were triggered per month. Furthermore, pressure relief holes in soft coal areas commonly experienced collapse, resulting in rapid attenuation of pressure relief effects.

[0065] The specific implementation process of the deep rockburst mine tunneling and anchoring synchronous pressure relief and hazard mitigation technology method of the present invention is as follows: The first step was advanced detection and target area calibration. An airborne ground-penetrating radar was mounted on the roadheader to detect the surrounding rock within a 20m radius before each cycle of operation. Combined with existing geological data, the changes in coal seam thickness, roof lithology distribution, and relative magnitude of ground stress were obtained. The detection showed that the coal seam thickness in this test section was stable, with no obvious large-scale structures, but localized minor stress fluctuations, and the overall impact hazard level remained moderate. Based on the detection results, graded stress relief target areas were calibrated: the coal seam side target area was the range from the roadway side surface to a depth of 10m, divided into a surface energy release zone, a middle damping zone, and a deep conduction zone according to the stress gradient; the roof stress relief target area was the rock strata range of 3-8m below the basic roof; and the floor stress relief target area was the range from the floor surface to a depth of 4m. Two calibration holes were constructed before construction to complete the localization calibration of the frictional heat generation benchmark and stress inversion model. The first threshold for the stress concentration factor was set to 1.6, and the second threshold to 1.2.

[0066] The second step involves simultaneous and parallel excavation and unloading operations. A single-cycle advance of 2m is set, and the cutting unit of the integrated excavator and anchor machine performs cutting operations according to the rectangular roadway cross-section, with a single-cycle cutting time of approximately 26 minutes. Simultaneously with the start of the cutting operation, two drilling rigs on the left and right sides of the machine simultaneously construct gradient energy-dissipating and pressure-relief holes on both sides of the roadway. Four pressure-relief holes are constructed per side per cycle, with an initial hole spacing of 1.5m. The holes are horizontally arranged and perpendicular to the roadway sides. The parameters for the gradient energy-dissipating and pressure-relief holes are set as follows: original hole diameter 94mm, hole depth 10m; hole opening diameter 150mm, length 2m; middle necking section diameter 70mm, length 1.5m, located at a depth of 2-3.5m below the hole opening, forming two annular support steps at the diameter change points. The drilling process employs a method of first drilling the original hole and then enlarging it in sections. The construction time for a single hole is approximately 6.5 minutes. Eight holes on both sides are drilled in parallel, with a total construction time of approximately 26 minutes. This process runs concurrently with the cutting operation and does not require additional cycle time.

[0067] After drilling is completed, the energy dissipating body at the bottom of the borehole, the rigid partition, the necking section energy dissipating body, the rigid partition, and the borehole opening section energy dissipating body are sequentially inserted using the push rod provided with the drilling rig. Finally, an embedded limiting retaining ring is driven into the borehole opening. The energy dissipating body at the borehole opening has a diameter of 148mm, a yield strength corresponding to a stress of 8MPa, and a pre-fabricated deep groove on its end face; the necking section has a diameter of 68mm, a yield strength corresponding to a stress of 12MPa; and the bottom section has a diameter of 92mm, a yield strength corresponding to a stress of 18MPa. The flange of the energy dissipating body precisely abuts against the corresponding step surface, with no axial movement or gaps. After insertion, it provides continuous support to the borehole wall, preventing borehole collapse in soft coal areas. When the surrounding rock stress increases, the energy dissipating body at the borehole opening first yields and dissipates energy along the induction groove. As the stress continues to rise, the middle and deep sections are activated sequentially, achieving a graded and gradual energy release.

[0068] Simultaneously, the airborne top-mounted advanced drilling rig drills biomimetic dendritic directional cutting holes in the roof slab ahead of the drilling face. An inclination sensor built into the drill rod corrects the trajectory in real time. The drilling elevation angle is set at 15°, the hole depth at 15m, and it maintains a distance of 8m ahead of the drilling face at all times. After drilling to the designed depth, a pulsed high-pressure water jet device is connected. First, a main seam is formed along the fine sandstone layer, with a length of 12m and a cutting height of approximately 0.5m. Then, the nozzle is deflected by 30°, and a branch seam is cut every 2m along the main seam axis, with a branch seam length of 3m. During the cutting process, the water pressure is adjusted in real time based on trajectory feedback. When encountering bedding planes, the feed rate is reduced to ensure that the angle and length of the branch seams meet the design requirements. After the seams are cut, the high-pressure air-carrying sand mode is switched, injecting high-strength ceramic particles with a particle size of 0.5~1mm into the seams. The injection volume per hole is approximately 8kg, ensuring that the ceramic particles are evenly supported between the fracture surfaces to prevent the seams from closing. The top plate cutting operation is carried out simultaneously with the cutting and side pressure relief operation. The hole changing and splicing are completed by utilizing the circulation gap, without increasing the single cycle time.

[0069] The third step involves multi-source coupled stress sensing and parameter control. During the construction of the pressure relief holes in the sidewall, drilling torque and propulsion resistance are collected in real time through armored sensor groups at the front end of the drill pipe. Simultaneously, the temperature of the drill cuttings at the bottom of the hole is collected through armored platinum resistance thermometers, with a sampling frequency of once per second. At the same time, micro-fracture signals of the coal seam are collected through miniature acoustic emission sensors pre-positioned in the sidewall.

[0070] Before the inversion calculation, a double correction was performed: based on the real-time collected drilling pressure of 12kN, rotation speed of 120r / min, and flushing fluid flow rate of 40L / min, the corresponding frictional heat value of 1.8℃ was retrieved from the calibration reference library and subtracted from the original temperature measurement value; then, based on the 10m hole depth and the geothermal gradient of 2.8℃ / 100m in the mine, the geothermal increment of 0.28℃ was subtracted, finally obtaining the stress-induced temperature increment. The drilling torque, the corrected temperature increment, and the number of acoustic emission frequencies were substituted into the localized mapping model with weights of 0.6:0.2:0.2 to calculate the stress concentration factor at the current location.

[0071] During this construction cycle, the measured torque in the normal drilling section was 120~150 N·m, the corrected temperature increment was 0.7℃, the average frequency of acoustic emission events was 7 times / min, and the calculated stress concentration factor was 1.35, which is within the normal range of 1.2~1.6. Therefore, the initial pressure relief hole parameters and energy dissipation body level were maintained unchanged. When drilling the No. 3 sidewall borehole, the measured drilling torque increased to 175 N·m, the corrected temperature increment increased to 1.6℃, and the acoustic emission frequency increased to 11 times / min. The inverted stress concentration factor was 1.68, exceeding the first threshold of 1.6, and was identified as a local high-stress area. Subsequently, the construction parameters of subsequent boreholes were adjusted, the spacing of pressure relief holes in this area was reduced to 1.2m, the length of the enlarged section was increased to 2.5m, and a higher-level phase change energy dissipation body was used. At the same time, the branch density of the top plate cut was increased to improve the local pressure relief intensity and ensure the pressure relief effect in the high-stress area.

[0072] The fourth step involves staggering the timing of support and unloading operations and relieving pressure on the bottom plate. After the single-cycle cutting and pressure relief operations on the side and top plate are completed, the cutting unit is withdrawn, and the temporary initial support operation is immediately started. The top and side are temporarily supported by quick-installing end anchor bolts. The installation time for each anchor bolt is about 80 seconds, and a total of 8 temporary anchor bolts are installed, taking about 11 minutes. This quickly forms temporary constraints on the surrounding rock and prevents the side from fracturing during subsequent operations.

[0073] After the initial support was completed, the permanent support anchoring operation and the static pressure relief operation of the floor slab were started simultaneously. The permanent support construction included top anchor bolts, side anchor bolts, and roof anchor cables. The side anchor bolts were 2.4m long threaded steel bolts with an anchoring section length of 1.2m. After installation, the bottom of the anchoring section was located 2.4m below the surface of the roadway sidewall, precisely within the original borehole section area of ​​the gradient energy-dissipating pressure relief hole. This completely avoided the enlarged borehole section and phase change energy-dissipating body within 2m of the borehole opening, ensuring the anchor bolts were anchored within the intact coal seam. The top anchor bolt spacing was 0.8m, and the side anchor bolt spacing was 1.0m. The total permanent support time was approximately 22 minutes.

[0074] The simultaneous floor pressure relief work employed a hydraulic static fracturing technique. Each cycle involved drilling six short pressure relief holes, 4m deep and spaced 2m apart, arranged in a quincunx pattern. After drilling, a hydraulic fracturing device was inserted, and a high-pressure pump station progressively increased pressure to induce directional fractures in the rock mass, forming a floor loosening and pressure relief zone approximately 4m deep. This process involved no blasting vibrations or impact loads, ensuring no disturbance to the curing anchoring agent and guaranteeing support quality. The floor loosening zone connected with the gradient energy-dissipating pressure relief holes on both sides at the bottom of the space, forming a continuous pressure relief ring around the tunnel, transferring high stress from the surrounding area to deeper layers.

[0075] Step 5: Lag Verification and Cyclic Advancement. After a single cycle of operation is completed, the tunneling and anchoring machine advances 2 meters to prepare for the next cycle. Simultaneously, at a position 10 meters behind the face, the stress relief effect is verified using a combination of drill cuttings method, online stress monitoring, and borehole imaging: two boreholes are inspected on each face, and the maximum drill cuttings amount is measured to be 3.1 kg / m, which is lower than the critical value of 4.5 kg / m set for the mine; online stress monitoring data shows that the coal stress shows a continuous downward trend; borehole imaging shows that the borehole wall is intact, with no collapse, and the development of stress relief fractures is within the design range. Based on the comprehensive assessment, the stress relief effect is deemed satisfactory, and the machine proceeds normally to the next cycle.

[0076] This embodiment underwent continuous testing for 30 days, with a cumulative advance of 275m. The monthly advance can reach over 270m, which is about 52% higher than the original process. The average approach distance between the two sides of the roadway was 132mm, which is 37.1% lower than the original process. The anchor bolt anchoring force qualification rate reached 99.3%, and there were no cases of insufficient anchoring force. During the test, only one minor impact hazard warning was triggered, and there was no collapse of the pressure relief hole. The effective action period of the cutting was extended by more than 2 times compared with the traditional process, and the overall prevention and control effect was significant.

[0077] Example 2: This embodiment was applied in a deep, thick, hard-roofed mine with a high risk of rockburst. The mine has a mining depth of 1120m, a main coal seam thickness of 4.1m, a uniaxial compressive strength of 16MPa, and a roof of 14m thick, hard sandstone with a uniaxial compressive strength of 68MPa. It is a typical rockburst mine with a thick, hard roof, a maximum horizontal principal stress of 28.5MPa, an initial stress concentration factor of 1.8, and a rockburst hazard level of strong impact. The mine originally used a process of deepening the borehole and blasting to relieve pressure. The pressure relief operation took up a separate working time, with a single cycle time of 70 minutes and a monthly advance of only 125m. Moreover, the roof pressure was strong, and strong coal blasting occurred frequently. The blasting vibration also often led to a decrease in the anchoring force of the surrounding anchor bolts, resulting in high construction safety risks.

[0078] The process method of this invention optimizes parameters to suit the characteristics of high impact and thick, hard top plates. The specific implementation process is as follows: The first step was advanced detection and target area calibration. Airborne ground-penetrating radar combined with advanced borehole stress gauges was used to detect the surrounding rock structure and stress distribution within a 30m radius ahead, focusing on identifying the strata, thickness, and primary fracture distribution of the thick, hard sandstone roof. Three calibration boreholes were drilled before construction to complete the calibration of the frictional heat generation benchmark and inversion model under high stress. Detection revealed that the sandstone thickness in the test section was stable, with no obvious fault structures, but the overall stress level was high, and the roof strata showed good integrity and strong energy accumulation capacity. The graded stress relief target areas were calibrated: the coal seam side target area was 12m deep, divided into three energy release gradients; the roof stress relief target area was the middle and lower strata of the thick, hard sandstone, 5-12m above the roadway roof, divided into two fractured target areas; the floor stress relief target area was 5m deep. The first threshold for stress concentration factors was set at 1.7, and the second threshold at 1.4, indicating that the overall stress relief intensity was higher than that of medium-impact mines.

[0079] The second step involves simultaneous excavation and unloading operations. A single-cycle advance is set at 1.8m, with a single-cycle cutting time of approximately 30 minutes. Simultaneously, gradient energy-dissipating and pressure-relief holes are constructed in the sidewalls and biomimetic dendritic cutting holes in the roof. The initial spacing between the sidewall pressure-relief holes is set at 1.2m, with 5 holes constructed per sidewall per cycle. The gradient hole parameters are: original hole diameter 94mm, hole depth 12m; enlarged hole diameter 180mm, length 2.5m; necking hole diameter 75mm, length 2m, located 2.5~4.5m below the hole opening, with two variable-diameter steps supporting the corresponding energy-dissipating body. The accompanying phase change energy-dissipating body is more finely graded. The hole opening section uses a low-yield-strength, high-energy-dissipating formula to enhance surface energy dissipation. The necking and bottom sections progressively increase the yield threshold. Each section's end face has a pre-fabricated induction groove of corresponding depth, combined with rigid spacers to ensure stable graded yielding sequence under high stress, avoiding instantaneous energy release that could induce impact. The borehole opening is equipped with a widened limiting retaining ring to accommodate axial thrust under high stress. The side drilling rig operates in parallel, with a total duration of approximately 29 minutes, and is completed simultaneously with the cutting operation.

[0080] For thick, hard sandstone strata, the pressure relief effect of biomimetic dendritic cutting was enhanced. A double-row arrangement of cutting holes was used, with a 2m spacing between upper and lower layers, corresponding to different strata of the sandstone. The drilling angles were 12° and 18° respectively, with a hole depth of 20m and an advance head distance of no less than 10m. The drill rod trajectory was monitored throughout the entire process. Layered dendritic cutting was performed using ultra-high pressure pulsed water jets. The main fracture length per hole was 15m, with a cutting height of 0.6m per layer. Branch fractures were densely arranged, spaced 1.5m apart, with a length of 4m. The angle was corrected in real time during cutting to avoid deviation along the original fractures. After cutting, ceramic proppant was injected twice to ensure effective support for both the branch and main fractures. This formed two long-lasting, three-dimensional dendritic fracture networks within the thick, hard sandstone, cutting the thick, hard rock layer into a multi-layered substructure. This significantly reduced its energy accumulation capacity and fracture impact strength, while preventing disordered fractures from causing overall instability of the strata. The roof slab cutting operation utilizes multiple drilling rigs in parallel, and is carried out simultaneously with the cutting and sidewall decompression, achieving zero additional time consumption through process integration.

[0081] The third step involves multi-source coupled stress sensing and parameter control. To address the risk of strong impacts, the spacing between acoustic emission sensors was increased, and the acquisition frequency of drill cuttings temperature and drilling load parameters was raised. A dual-calibration model was used to ensure inversion accuracy. During construction, the stress concentration factor was inverted in real time. When the factor exceeded the first threshold of 1.7, the spacing between pressure relief holes was reduced to 1.0m, the length of the enlarged borehole section was increased to 3m, and a higher-grade energy dissipator was used. Simultaneously, the branch density of the top plate cuts and the amount of ceramsite injected were increased to enhance the pre-cracking effect of the top plate. In this embodiment, the overall stress level in the test section was relatively high, with the stress concentration factor obtained from inversion in most areas between 1.6 and 1.8. Therefore, the reduced hole spacing and higher-grade energy dissipator were generally used, and drilling was further intensified in local areas of abnormal stress. Through dynamic control, the pressure relief strength in high-stress areas was ensured while excessive construction in low-stress areas was avoided.

[0082] The fourth step involves staggered support and decompression of the floor. After cutting and decompression, high-strength temporary initial supports are quickly constructed, and temporary anchor bolts are densely arranged to form a stable temporary load-bearing ring. Subsequently, permanent support and static cracking decompression of the floor are carried out simultaneously. The permanent support adopts a combined support system of high-strength anchor bolts and anchor cables. The length of the side anchor bolts is 2.8m, and the anchoring section is 1.5m long. The entire anchoring section is located in the intact coal seam below the gradient hole necking section to ensure the reliability of anchoring under high stress. The roof is supported by long anchor cables, with the anchor cable length extending into the stable rock layer above the dendritic cut, forming a cooperative load-bearing structure with the pre-fractured roof, which allows for rock deformation and energy release while effectively controlling the amount of roof subsidence.

[0083] The base plate pressure relief employs deepened hydraulic static cracking holes, 5m deep and spaced 1.8m apart, with the expansion pressure increased in stages to create a deeper loosening zone in the base plate, thus alleviating stress concentration. The entire static cracking operation of the base plate is vibration-free and runs completely in parallel with the permanent support. The total time for a single cycle of support and base plate pressure relief is approximately 28 minutes, which is basically matched with the cutting time.

[0084] The fifth step is lag verification and cyclical advancement. Decompression verification is conducted at a position 12m behind the face, using a combination of drill cuttings method, microseismic monitoring, and roof delamination monitoring to comprehensively assess the decompression effect. During the test, the amount of drill cuttings was controlled below the critical value, the frequency of high-energy microseismic events decreased by 79% compared to the original process, no strong coal blasting occurred, and the amount of roof delamination was controlled within a safe range. Subsequent spot checks showed that the roof cuts maintained good conductivity and no large-area closure occurred.

[0085] After the implementation of this embodiment, the single-cycle operation time is controlled within 33 minutes, the monthly advance is increased to 218m, which is 74.4% higher than the original process; the roadway roof subsidence is reduced by 45%, the sidewall movement is reduced by 40%, the impact hazard events are greatly reduced, and the anchor bolt anchoring force qualification rate is increased to 98.9%, realizing safe and efficient tunneling in mines with strong impact and thick hard roof.

[0086] Example 3: This embodiment was applied to a roadway with a high risk of impact near a fault structure. The roadway excavation work faced a normal fault with a drop of nearly 5m. Due to the influence of tectonic stress, the stress distribution was extremely uneven, with the local stress concentration factor exceeding 2.0, and the impact hazard level was strong impact. The original process used large-diameter boreholes with static parameters for pressure relief, which frequently resulted in insufficient or excessive pressure relief in some areas, triggering high-level impact warnings multiple times. In addition, the coal seam in the structural zone was fractured, the pressure relief boreholes collapsed severely, and the excavation speed was slow, with a monthly advance of less than 100m.

[0087] The process method of this invention focuses on leveraging the advantages of multi-source dynamic control and phase-change wall protection, adapting to the characteristics of large stress fluctuations and coal body fragmentation in structural zones. The specific implementation process is as follows: The first step was advanced detection and target area calibration. The frequency of advanced detection was increased, with ground-penetrating radar detection and advanced borehole stress testing conducted in each cycle to accurately identify fault locations, fracture zone ranges, and areas of tectonic stress concentration. Given the significant lithological variations within the tectonic zone, calibration tests were conducted in segments to establish frictional heat generation benchmarks and inversion parameters under different coal and rock strengths. Detection revealed that the fault was located 25m ahead of the face, with a fault-affected zone approximately 15m wide. Within this zone, the coal was fractured, stress was significantly increased, and the stress distribution was extremely uneven. The calibration target area within the fault-affected zone was expanded, and the stress-relief intensity gradient was denser. Conventional parameters were used for the non-affected zone. Three stress threshold levels were set, and the stress-relief level was dynamically switched according to the stress level to achieve refined stress relief within the tectonic zone.

[0088] The second step involves simultaneous and parallel excavation and unloading operations. A single cycle advances 1.5m, with a cutting time of approximately 22 minutes. Gradient stress relief holes in the sidewalls and branched slotted holes in the roof are constructed concurrently. In normal areas far from the fault, moderate impact stress relief parameters are used. As excavation approaches the fault, the hole spacing is gradually reduced, the enlarged hole length is increased, and the energy dissipation body level is upgraded based on the detected stress increase trend. The sidewall stress relief holes utilize rapidly changeable enlarged drilling tools, allowing for real-time adjustment of the enlarged hole diameter and length during construction. Phase change energy dissipation bodies are pre-classified and stored according to their levels, allowing for rapid replacement based on instructions to adapt to rapid stress changes. Energy dissipation bodies are immediately implanted after drilling within the structural zone. The continuous support of these bodies prevents the collapse of the broken coal seam, effectively solving the problem of easy clogging of stress relief holes in structural zones.

[0089] The top plate slots are densely arranged within the fault influence zone to increase the depth of the main slots and the density of branches. At the same time, the slot angle is adjusted to avoid the core area of ​​the fault fracture zone to prevent the slots from conducting through the fault and causing instability of the surrounding rock. After the slots are completed, the amount of ceramsite injected is increased to ensure the long-term conduction of the slots under tectonic stress.

[0090] The third step involves multi-source coupled stress sensing and parameter control. Addressing the characteristics of abrupt stress changes and variable lithology in the tectonic zone, the multi-source coupled inversion model is optimized to improve response speed, enabling stress inversion to be performed every 1 meter of borehole drilling. During the inversion process, correction parameters for corresponding segments are called in real time based on lithological changes to ensure inversion accuracy under different coal and rock strengths. When the drill load suddenly fluctuates, the corrected temperature rises sharply, or the acoustic emission frequency increases abruptly during drilling, it is immediately identified as a stress anomaly zone. The borehole enlargement parameters and energy dissipation level are adjusted on-site, and the density of subsequent boreholes is increased. During construction within the fault influence zone, multiple instances of abrupt stress changes occurred; in each case, precise stress relief was achieved through dynamic parameter adjustment, preventing localized stress accumulation.

[0091] The fourth step involves staggered support and unloading sequences and pressure relief of the base slab. Support operations utilize yielding anchor bolts in conjunction with high-strength anchor cables, along with gradual energy release through gradient pressure relief holes, to achieve deformation coordination between support and pressure relief. Temporary initial supports are densely arranged within the structural zone to enhance temporary bearing capacity and prevent roof collapse during construction. Pressure relief of the base slab is achieved by deepening the holes to 6m within the fault influence zone, strengthening pressure relief and mitigating the concentration of structural stress in the base slab. All procedures are carried out synchronously and in parallel, without increasing cycle time due to the increased pressure relief intensity.

[0092] The fifth step is delayed verification and cyclical advancement. The frequency of pressure relief verification within the fault influence zone is increased, with a comprehensive verification conducted every 5 meters to ensure the pressure relief effect meets standards and the surrounding rock remains stable. After safely passing through the fault influence zone, the pressure relief parameters are gradually adjusted back to normal levels.

[0093] This embodiment successfully and safely traversed the fault zone, with no strong impact warnings or pressure relief hole collapses during tunneling. The monthly advance reached 188m, an improvement of over 88% compared to the original process, verifying the applicability and reliability of this process in complex structural areas.

[0094] Compare with Example 1: Using the same mine geological conditions as in Example 1, and employing the traditional "first excavate and anchor, then depressurize" process, conventional large-diameter depressurization holes with a diameter of 150mm, a depth of 10m, and a spacing of 1.5m were used. Depressurization was carried out separately after the excavation and anchoring operations were completed, without phase change energy-consuming bodies or roof directional cutting. Testing showed a total single-cycle time of 58 minutes and a monthly advance of approximately 150m, significantly lower than in Example 1; the roadway sidewall convergence was 245mm, the anchor bolt anchoring strength qualification rate was 82%, and the hole collapse rate in the soft coal area reached 32%, indicating a rapid decline in depressurization effect; the number of impact warnings was 8.5 per month, and the overall prevention and control effect was far inferior to the process of this invention.

[0095] Compare with Example 2: Using the same synchronous operation mode as in Example 1, the pressure relief holes adopted a conventional equal-diameter structure, without phase change energy-consuming bodies or positioning structures. The roof was filled with ordinary horizontal cuts without proppant. Temperature parameters were directly coupled without correction and without dynamic parameter control. Fixed hole spacing and diameter were used, and conventional blasting pressure relief was employed for the floor. Testing showed that although the tunneling efficiency was similar to Example 1, the coal seam at the borehole opening was severely fractured, the anchor bolt anchoring force qualification rate was only 84%, and the floor blasting caused a decrease in anchor bolt anchoring force of approximately 15%. The stress inversion error reached 23%, and insufficient pressure relief occurred in localized stress anomaly zones, with 5.2 warnings per month. The roof cut closure rate reached 65% after 30 days, and the pressure relief effect continued to decline. Overall, the performance was far inferior to the process of this invention.

[0096] The above description of the disclosed embodiments enables those skilled in the art to make or use the invention. Various modifications to these embodiments will be readily apparent to those skilled in the art, and the general principles defined herein may be implemented in other embodiments without departing from the spirit or scope of the invention. Therefore, the invention is not to be limited to the embodiments shown herein, but is to be accorded the widest scope consistent with the principles and novel features disclosed herein.

Claims

1. A method for simultaneous depressurization and hazard mitigation in deep rockburst mines, characterized in that, Includes the following steps: S1: Before tunneling, detect the geological parameters and stress distribution of the surrounding rock ahead, classify the impact hazard level, mark the graded stress relief target areas corresponding to the roof, coal seam and floor respectively, and set stress control intervals including the first threshold and the second threshold for each target area. S2: A single-cycle operation is performed by a tunneling and anchoring machine. While the cutting section is cutting the roadway, a machine-mounted side drilling rig is used to construct gradient energy-dissipating and pressure-relieving holes on both sides of the roadway and implant phase change energy-dissipating bodies. A machine-mounted top-advanced drilling rig is used to construct biomimetic dendritic directional cutting holes in the roof in front of the face. The gradient energy-dissipating and pressure-relieving holes are axially variable diameter structures, and phase change energy-dissipating bodies with yield thresholds gradually increasing from shallow to deep are set in sections inside to form graded progressive energy-dissipating channels. The biomimetic dendritic directional cutting holes are formed by water jet cutting to create main and branch seams and injecting supporting particles to construct a directional fracture pressure-relieving network. S3: During the pressure relief drilling process, drilling response parameters and surrounding rock response signals are collected. After interference correction, the stress concentration coefficient of the surrounding rock in the current construction area is obtained by the stress concentration coefficient inversion method. The inverted stress concentration coefficient is compared with the stress control interval. When the stress concentration coefficient is higher than the first threshold or lower than the second threshold, the construction parameters of the subsequent pressure relief holes and slotted holes that have not yet been constructed in this cycle are adjusted accordingly. S4: After the single-cycle cutting and pressure relief of the side and top plate are completed, temporary initial support is constructed first to maintain the stability of the surrounding rock, and then permanent support anchoring and static directional rock-cracking pressure relief of the bottom plate are carried out simultaneously; the anchoring section of the permanent anchor rod avoids the borehole expansion section of the gradient energy dissipation pressure relief hole. S5: After completing a single cycle of pressure relief and support, the tunneling and anchoring machine advances one cycle of progress. The pressure relief effect is checked at a set distance behind the face, and the check result is used as the basis for correction of the target area calibration of the next cycle of graded pressure relief. After the check is qualified, the next working cycle begins.

2. The method for simultaneous depressurization and hazard mitigation in deep rockburst mines according to claim 1, characterized in that, The gradient energy dissipation and pressure relief hole is provided with an enlarged orifice section, a narrowed middle section, and an original orifice section in sequence along the axial direction; the diameter of the enlarged orifice section is 1.5 to 2 times the diameter of the original orifice section, and the length is 1.5 to 3 m; the diameter of the narrowed middle section is 0.6 to 0.8 times the diameter of the original orifice section, and the length is 1 to 2 m; the junctions of the enlarged orifice section and the narrowed middle section, and the junctions of the narrowed middle section and the original orifice section form an annular support step.

3. The method for simultaneous decompression and hazard relief in deep rockburst mines according to claim 2, characterized in that, An embedded limiting ring is provided at the opening of the gradient energy dissipation and pressure relief hole. The end of the phase change energy dissipator has an annular flange, which abuts against the annular support step and the retaining ring to achieve axial fixation. An annular deformation induction groove is prefabricated on the end face of the energy dissipator. The yield threshold is gradually increased from shallow to deep. When under pressure, the energy is dissipated in sequence according to the hole depth.

4. The method for simultaneous depressurization and hazard mitigation in deep rockburst mines according to claim 1, characterized in that, The biomimetic dendritic directional slotting borehole has a lead distance of no less than 5m and an elevation angle of 10°~20°. During the drilling process, the drill rod has a built-in tilt sensor to correct the drilling trajectory in real time. After drilling to the design depth, a pulsed high-pressure water jet is used to cut along the key layer of the top plate to form the main slot. The length of the main slot in a single hole is no less than 10m. Then, multiple branch slots are cut along the axis of the main slot at an angle of 25°~35°. After the slotting is completed, high-pressure air is used as a transport carrier to carry the supporting particles into the fissure. The supporting particles are high-strength ceramic particles that support the fissure surface to resist the closure of the confining pressure.

5. The method for simultaneous depressurization and hazard mitigation in deep rockburst mines according to claim 1, characterized in that, The inversion method for the stress concentration factor includes: A baseline database corresponding to drilling pressure, rotation speed, flushing fluid flow rate, and frictional heat generation is established in advance through calibration holes on the working face. Drilling parameters are collected in real time and the frictional heat generation component is deducted. Then, the geothermal gradient is corrected according to the borehole depth to obtain the stress-induced temperature increment. Drilling torque, corrected temperature increment, and acoustic emission event frequency are coupled according to weights to establish a mapping model to calculate the stress concentration factor of the surrounding rock at the current location. When the stress concentration factor is higher than the first threshold, the spacing of the pressure relief holes is reduced, the length of the enlarged section is increased, and the energy consumption level of the energy-consuming body is improved, while the cutting branch density is increased. When the stress concentration factor is lower than the second threshold, the spacing of the pressure relief holes is increased and the length of the enlarged section is shortened, while the cutting branch density is reduced. The first threshold is greater than the second threshold.

6. The method for simultaneous decompression and hazard relief in deep rockburst mines according to claim 1, characterized in that, The acquisition of drilling response parameters and surrounding rock response signals includes: The drilling torque and propulsion resistance are collected in real time by the armored sensor group built into the front end of the drill pipe, the temperature of the drill cuttings at the bottom of the hole is directly collected by the armored platinum resistance temperature sensor, and the micro-fracture signal of the surrounding rock is collected simultaneously by the miniature acoustic emission sensor in the roadway side; at the same time, the drilling pressure, rotation speed and flushing fluid flow rate are collected. The interference correction includes frictional heat generation correction and geothermal gradient correction; The frictional heat generation correction is as follows: a reference library corresponding to drilling pressure, rotation speed, flushing fluid flow rate and drill bit frictional heat generation is established in advance through the calibration hole of the working face; during construction, drilling pressure, rotation speed and flushing fluid flow rate are collected in real time and the corresponding frictional heat generation is deducted. The geothermal gradient correction is performed by subtracting the original geothermal temperature increment with depth based on the borehole depth and the regional geothermal gradient value.

7. The method for simultaneous depressurization and hazard mitigation in deep rockburst mines according to claim 1, characterized in that, The temporary initial support adopts quick-installation end anchor bolts, with a single bolt installation time not exceeding 90 seconds. The pressure relief operation of the base plate is started immediately after the initial support is completed. The permanent support anchoring and the static directional rock-cracking pressure relief of the base plate are carried out in parallel. The pressure relief short holes are constructed on the base plate using hydraulic static cracking technology. The rock-cracking depth of the pressure relief short holes is 3~6m, and the hole spacing is 1.5~2.5m. The resulting loosening zone of the base plate is spatially connected with the gradient energy-dissipating pressure relief holes on both sides.

8. The method for simultaneous depressurization and hazard mitigation in deep rockburst mines according to claim 1, characterized in that, Step S2 includes: The parallel overlap of the three actions performed by the cutting department—cutting the roadway, constructing gradient energy-consuming and pressure-relieving holes on both sides of the roadway using a machine-mounted side drilling rig, and constructing biomimetic dendritic directional cutting holes in the roof of the face using a machine-mounted top advanced drilling rig—is no less than 80%, and the total duration of these three actions does not exceed 1.15 times the cutting operation duration.

9. The method for simultaneous depressurization and hazard mitigation in deep rockburst mines according to claim 1, characterized in that, The pressure relief effect is verified by a combination of drill cuttings method, online stress monitoring and borehole wall imaging. When the amount of drill cuttings is lower than the critical value, the stress monitoring data shows a downward trend and the development of borehole wall cracks is within the design range, the pressure relief is deemed qualified.

10. The method for simultaneous depressurization and hazard mitigation in deep rockburst mines according to claim 1, characterized in that, The advance of one cycle is 1.5~2.5m, the pressure relief and verification position is located 8~15m behind the face, and the biomimetic dendritic directional cutting borehole always maintains a lead distance of not less than 3m.

Images