Working face end top plate weakening and scour preventing method

Abstract

The invention discloses a working face end head top plate weakening anti-scour method, which adopts a mode of performing hydraulic slotting in a surrounding rock top plate of a roadway dangerous area of a working face end roadway to form a pressure relief slot, a retarding slot and a far-field dynamic load source breaking and expanding slot so as to improve the anti-scour capability of the working face end roadway. A pressure relief slot formed by hydraulic slotting on a roadway top plate can release static load of roadway surrounding rock, and a retarding slot formed by hydraulic slotting on a dynamic load propagation path can retard far-field dynamic load from being propagated to the roadway. The far-field dynamic load source breaking and expanding seam groove formed by hydraulic slotting in the rock mass near the dynamic load source can reduce the elastic energy density of the rock mass near the dynamic load source and reduce the possibility of roof breaking. According to the method, the energy capable of inducing the rock burst disaster can be jointly reduced in the three aspects of the generation condition of the dynamic load, the propagation path of the dynamic load and the static load of the surrounding rock, and the method is particularly suitable for solving the rock burst problem of the end roadway of the working face.

Description

Technical Field

[0001] The present invention relates to a method for weakening and preventing impact of a roof plate at the end of a working face, specifically a method for weakening and preventing impact of a roof plate at the end of a working face, which can reduce the risk of impact mining pressure in the tunnel from three aspects: the generation conditions of dynamic loads, the propagation path of dynamic loads and the static load of surrounding rocks, and belongs to the technical field of safe mining of coal mines. Background Art

[0002] Coal mining in my country is mainly carried out in underground mining. With the increasing depletion of shallow resources, coal mining has gradually expanded to deep underground, which has led to increased stress in the coal seams, increased mining disturbance, and intensified strong mine pressure faced by coal mine production. A large number of studies have been conducted at home and abroad on the manifestation law of hard roof mine pressure and its disaster-causing mechanism. Rock burst is an instability phenomenon induced when the energy released by the coal-rock system is greater than the energy consumed. Rock burst will not occur when the stress on the coal-rock mass does not exceed the ultimate strength, and when the rate of elastic energy release in the coal-rock system is less than the rate of energy dissipation during crack expansion. However, mining engineering activities will lead to the readjustment of the stress and energy of the surrounding rock in the mining space. When the static load of the surrounding rock and the dynamic load caused by the external dynamic load are superimposed and exceed the critical load of coal-rock impact failure, the coal-rock system will be unstable and damaged, and rock burst will be induced. Especially when the roof of the working face is a hard roof, as the working face advances, the suspended roof above the goaf is too long, and the roof at the end of the working face is prone to accumulate a large amount of elastic energy. If a sudden fracture occurs under the action of mining stress or impact loads generated by geological activities, it is possible to induce a roof accident in the tunnel and then induce impact mine pressure.

[0003] In the prior art, the preventive measures against impact mining pressure are mainly divided into four solutions, namely, optimizing tunnel layout, strengthening support, filling and anti-impact, and weakening the hard roof. Among them, in terms of optimizing tunnel layout, the stress reduction zone in the goaf can be effectively utilized to weaken the static load effect on the tunnel; in terms of strengthening support, the deformation of the tunnel surrounding rock can be effectively controlled by using high prestressed anchor rods. However, neither of these two methods can fundamentally solve the high stress area of ​​the roof caused by the superposition of dynamic and static loads. Although filling and anti-impact can effectively solve the problem of suspended roof, the cost is high. Therefore, the current preventive measures against impact mining pressure mainly focus on weakening the strength of the hard roof. Traditional roof weakening methods include water injection to weaken the roof, drilling pressure relief, hydraulic fracturing, drilling and blasting pre-cracking and other means. Among them, water injection to weaken the roof has great limitations, and is not effective for rock masses with small porosity, weak water absorption and incomplete fracture development; drilling unloading can reduce the stress concentration of the surrounding coal rock mass, but the impact range of a single borehole is limited; hydraulic fracturing cannot effectively evaluate the fracturing effect for roofs with complex structures and lithologies; drilling and blasting pre-cracking has a large unloading range, but the operation is difficult, and the dynamic load generated may induce impact mine pressure. It is difficult to apply on a large scale due to restrictions on safe operations and environmental protection. Summary of the invention

[0004] In response to the problems existing in the above-mentioned prior art, the present invention provides a method for weakening the roof of the working face end to prevent impact, which can jointly reduce the energy that can induce impact mine pressure disasters in three aspects: the conditions for generating dynamic loads, the propagation path of dynamic loads, and the static load of surrounding rocks. Thereby, the load borne by the working face end tunnel is lower than the critical value that can induce impact mine pressure disasters, and the anti-impact capacity of the working face end tunnel is improved. It is particularly suitable for solving the impact mine pressure problem of the working face end tunnel.

[0005] To achieve the above purpose, the method for weakening the top plate of the working face end to prevent impact is to improve the anti-impact capability of the working face end roadway by hydraulically cutting the top plate near the dynamic load source, the dynamic load transmission path and the roadway top plate, and specifically includes the following steps:

[0006] Step 1: Obtain mine data, conduct on-site sampling, prepare standard samples for experiments, and obtain basic mechanical parameter data, internal porosity data, fracture distribution data, and optimal hydraulic rock breaking technical parameters of the target fully mechanized caving working face coal seam and its goaf.

[0007] Step 2, based on the mine data in Step 1, construct a load-deformation finite element model of the adjacent working face goaf-coal pillar-tunnel-coal seam to be mined structure, and perform numerical simulation of the target working face mining process. Use the periodic fracture and collapse data of the roof and top coal during mine production, the coal seam mining stress data, and the tunnel surrounding rock deformation and separation data to verify and optimize the load-deformation finite element model, and form a simulation model that meets the actual production conditions and monitoring data of the mine site. Based on the simulation model, perform numerical simulation of the working face mining of the subsequent mining plan, calculate the tunnel area with high static load and large stress concentration coefficient and the roof fracture area of ​​the adjacent goaf, and identify the dangerous area of ​​the tunnel with the danger of impact mine pressure;

[0008] Step 3: According to the coordinates of the dangerous area of ​​the tunnel obtained in Step 2 and the coordinates of the far-field dynamic load source obtained by on-site microseismic monitoring, the hydraulic slot weakening anti-scouring conditions that can meet the anti-scouring requirements of the tunnel are set;

[0009] Hydraulic cutting weakening and anti-scouring must meet the following conditions:

[0010]

[0011] Where: Ω is the rock volume element; W is the energy transmitted from the dynamic load to the surrounding rock of the tunnel; U t is the strain energy of the rock element; ψ is the potential energy of the rock element; t is time; x j is the coordinate; σ j is the coordinate x j Stress at location;

[0012] in,

[0013] ψ=U s +U b +U cr

[0014] Where: U s is the energy consumed in the compaction stage; U b The energy consumed by the development of secondary cracks, the extension of primary cracks, and the final fracture surface energy; U cr It is the energy released during the friction process of secondary crack fracture;

[0015] Volume element strain energy U t The energy W transmitted to the surrounding rock of the tunnel by the dynamic load satisfies the following conditions:

[0016]

[0017] Where: s is the static load of the tunnel surrounding rock; d is the dynamic load of the tunnel surrounding rock;

[0018] Initial static load of tunnel surrounding rock σ 0s With initial dynamic load σ 0d The following conditions must be met:

[0019]

[0020] Where: H n is the vertical height of coal strata at different burial depths; ρ n is the density of different rock layers; g is the acceleration of gravity; M is the mass of the collapsed rock layer; v is the collapse velocity of the rock layer; A is the impact area; λ is the Lame constant of the rock layer; w is the stress wave frequency of the impact load;

[0021] in,

[0022]

[0023] Where: c, k, G, u are the vertical damping coefficient, vertical spring coefficient, shear modulus, and Poisson's ratio of the rock layer respectively; r is the equivalent radius of the collapsed rock mass; ρ is the density of the collapsed rock mass;

[0024] When there are hydraulic cutting joints in the tunnel surrounding rock, the static load σ s and the dynamic load σ of the tunnel surrounding rock d Different attenuation will occur, and the attenuated values ​​meet the following conditions:

[0025]

[0026] Where: r0 is the radius of the slit; r x is the distance between the surrounding rock position and the center of the cutting seam; C is the stress wave attenuation coefficient; x0 is the distance between the cutting seam and the dynamic load source; x is the distance between the surrounding rock and the dynamic load source;

[0027] Step 4, based on the simulation model of Step 2 and the hydraulic cutting weakening anti-scouring conditions of Step 3, firstly, simulate drilling and simulate hydraulic cutting for the dangerous area of ​​the tunnel in the simulation model to form pressure relief slots, blocking slots and far-field dynamic load source broken expansion slots, wherein the pressure relief slots are located in the tunnel roof in the dangerous area of ​​the tunnel, the blocking slots are located on the line connecting the dangerous area of ​​the tunnel and the far-field dynamic load source, and the cutting surface of the blocking slots is perpendicular to the dynamic load propagation direction, the far-field dynamic load source broken expansion slots are in the same rock layer as the far-field dynamic load source in the vertical direction, and the far-field dynamic load source broken expansion slots are located near the far-field dynamic load source in the horizontal direction, then simulate the dynamic load generated by the roof breaking at the far-field dynamic load source position, and simultaneously carry out numerical simulation of the working face recovery of the subsequent mining plan, so as to obtain the optimal hydraulic cutting weakening anti-scouring technical parameters that can weaken the dynamic load and reduce the static load;

[0028] Step 5, according to the optimal hydraulic cutting weakening and anti-bumping technical parameters obtained in Step 4, a drilling rig is arranged in the target working face tunnel of the mine, and holes are drilled into the tunnel surrounding rock roof respectively, and the top rock is crushed by impacting the high-pressure water jet ejected from the jet nozzle at the end of the drill pipe and ejected along the radial direction of the drill pipe. Through the rotation and retreat of the drill pipe, pressure relief slots, blocking slots and far-field dynamic load source broken expansion slots are constructed, and so on, all pressure relief slots, blocking slots and far-field dynamic load source broken expansion slots are completed;

[0029] Step 6, based on the regional and real-time monitoring results of the tunnel surrounding rock in the target area monitored on site, determine the impact mine pressure hazard of the target tunnel surrounding rock before and after hydraulic cutting, evaluate and verify the anti-impact effect of hydraulic cutting, and repeat Step 4 for areas where the anti-impact effect does not meet the hydraulic cutting weakening anti-impact conditions of Step 3, adjust the optimal hydraulic cutting weakening anti-impact technical parameters obtained in Step 4, and perform the hydraulic cutting construction work of Step 5 again based on the adjusted and optimized optimal hydraulic cutting weakening anti-impact technical parameters until the anti-impact effect meets the hydraulic cutting weakening anti-impact conditions of Step 3.

[0030] Furthermore, when numerically simulating the mining process of the target working face in Step 2, firstly, the mine data including the geological conditions and mining conditions of the coal seam of the target fully-mechanized caving working face, as well as the basic mechanical parameter data of the coal body and surrounding rock of the coal seam and its goaf of the target fully-mechanized caving working face, the internal porosity data and the fracture distribution data are imported into the load-deformation finite element model; then, the vertical stress distribution is calculated to obtain the area with high elastic energy density of the roof; the grid of the area with high elastic energy density of the roof is encrypted; finally, the numerical simulation of the mining process of the target fully-mechanized caving working face is carried out to obtain the stress distribution, deformation characteristics and periodic collapse data of the roof and top coal of the target fully-mechanized caving working face and roadway.

[0031] Furthermore, when simulating hydraulic cutting in the dangerous area of ​​the tunnel in the simulation model in Step 4, the hydraulic cutting depth meets the following conditions:

[0032]

[0033] Where: p is the jet pressure; R is the nozzle radius; is the internal friction angle of the roof; τ is the shear strength of the roof rock; c is the cohesion of the roof rock.

[0034] Furthermore, when simulating hydraulic cutting to form a pressure relief groove in Step 4, simulated drilling is performed near the junction of the top coal and the coal pillar in the tunnel, and simulated hydraulic cutting is performed near the side of the tunnel. The downward extension range of the hydraulic cutting does not exceed the interface between the roof and the coal seam, and the upward extension range of the hydraulic cutting does not exceed the direct roof.

[0035] Furthermore, when simulating the hydraulic cutting to form a pressure relief slot in Step 4, after the roof is weakened, when the direct top settlement height is ΔH, the roof pressure in the cutting area is

[0036]

[0037] Where: E2 is the direct top elastic modulus; L R1 is the radius of the pressure relief slot; θ1 is the elevation angle of the pressure relief slot drill pipe; H2 is the height of the direct roof; q is the pressure of the direct roof on the coal pillar;

[0038] Among them, the pressure relief slotted drill pipe elevation angle θ1 satisfies the following conditions:

[0039] θ1=arctan[(h1-h0) / l0]

[0040] In the formula: h1 is the tunnel height; h0 is the drilling rig height; l0 is the distance between the drilling rig and the coal pillar.

[0041] Furthermore, when obtaining the optimal hydraulic slotting anti-collision technical parameters that can weaken the dynamic load and reduce the static load in Step 4, the dynamic load attenuation rate, the stress-strain change characteristics of the tunnel surrounding rock and the stress transfer characteristics of the coal pillar area under different hydraulic slotting parameters are obtained by grouping numerical simulation of different hydraulic slotting parameters, and the main control parameters of hydraulic slotting anti-collision that affect the dynamic load attenuation rate, the stress-strain change characteristics of the tunnel surrounding rock and the stress transfer characteristics of the coal pillar area are judged and obtained. After verification and model optimization combined with the field observation results of the mine, the optimal hydraulic slotting anti-collision technical parameters are formed.

[0042] Furthermore, when evaluating and verifying the anti-impact effect of hydraulic cutting in Step 6, microseismic monitoring is carried out in the tunnel after cutting, the frequency of the tunnel affected by mine earthquakes before and after cutting is compared, the energy level of mine earthquakes before and after cutting is compared, the deformation of the surrounding rock of the tunnel is measured, and the deformation of the surrounding rock in the cutting area and the uncut area is compared. The measured data, theoretical model calculations and numerical simulation calculations are compared and verified, and the effect of far-field expansion cutting on reducing the far-field roof elastic energy density to suppress the occurrence of mine earthquakes is evaluated, the effect of blocking cutting on weakening dynamic loads is evaluated, and the effect of unloading cutting on reducing the static load of the tunnel surrounding rock is evaluated.

[0043] Compared with the prior art, the present method for weakening and preventing impact of the top plate at the end of the working face adopts the method of hydraulically cutting in the surrounding rock top plate in the dangerous area of ​​the tunnel at the end of the working face to form pressure relief grooves, blocking grooves and far-field dynamic load source expansion grooves to improve the anti-impact capacity of the tunnel at the end of the working face: the pressure relief grooves formed by hydraulic cutting in the tunnel top plate can release the static load of the tunnel surrounding rock, increase the degree of crushing of the hard tunnel top plate, reduce the strength of the top plate to transfer stress to deep rock, and at the same time increase the energy dissipation during the propagation of far-field dynamic loads; the blocking grooves formed by hydraulic cutting on the dynamic load propagation path can reduce the strength of the top plate to transfer stress to deep rock and block the propagation of far-field dynamic loads to the tunnel; the far-field dynamic load source expansion grooves formed by hydraulic cutting in the rock mass near the dynamic load source can reduce the elastic energy density of the rock mass near the dynamic load source, thereby reducing the possibility of roof breakage. It is possible to jointly reduce the energy that can induce impact mine pressure disasters in three aspects: the generation conditions of dynamic loads, the propagation path of dynamic loads and the static load of surrounding rocks. The load borne by the end tunnel of the working face can be lower than the critical value that can induce impact mine pressure disasters, and the anti-impact capacity of the end tunnel of the working face can be improved, thereby significantly reducing the possibility of impact mine pressure disasters in the end tunnel of the working face. At the same time, the construction process of hydraulic cutting is relatively refined and the engineering volume is small. The water jet rock breaking technology itself is relatively mature, the process is safe and green and pollution-free, and the construction difficulty is low. It is particularly suitable for solving the impact mine pressure problem of the end tunnel of the working face. BRIEF DESCRIPTION OF THE DRAWINGS

[0044] Figure 1 It is a top view layout diagram of the hydraulic cavitation construction of the present invention;

[0045] Figure 2 Section view along the working face. DETAILED DESCRIPTION

[0046] This method of weakening and preventing impact of the roof at the end of the working face achieves the goal of reducing the energy that can induce impact mining pressure disasters in three aspects: the generation conditions of dynamic loads, the propagation path of dynamic loads and the static load of surrounding rocks by hydraulically cutting the roof near the dynamic load source, the dynamic load propagation path and the roadway roof. Specifically: hydraulic cutting is performed on the rock mass near the dynamic load source to reduce the elastic energy density of the rock mass near the dynamic load source and the possibility of roof breakage; hydraulic cutting is performed on the dynamic load propagation path to block the propagation of far-field dynamic loads to the roadway; hydraulic cutting is performed on the coal pillar roof to release the static load of the roadway surrounding rocks.

[0047] The present invention will be further described below by taking an isolated working face tunnel of a coal mine working face as an example in conjunction with the accompanying drawings.

[0048] The method for weakening and preventing impact of the top plate at the end of the working face specifically comprises the following steps:

[0049] Step 1: Obtain mine data, conduct on-site sampling, prepare standard samples for experiments, and obtain basic mechanical parameter data, internal porosity data, fracture distribution data, and optimal hydraulic rock breaking technical parameters of the target fully-mechanized caving working face coal seam and its goaf.

[0050] The average burial depth of the comprehensive top coal mining working face of this coal mine is 450m, the average coal seam inclination is 6°, the average coal seam thickness is 7m, the direct roof thickness is 15.5m, and there is a 15m wide coal pillar between the goaf and the working face. The working face return air lane is excavated along the coal floor, with a rectangular cross section, a height h1 of 3.0m, and a width l1 of 5.0m. The support form of the return air lane is full anchor cable support and single hydraulic support. According to the monitoring data analysis, the length of the goaf roof during the target working face mining period is between 15 and 35m, and the working face periodic pressure step is 20m.

[0051] In order to accurately obtain the basic mechanical parameter data, internal porosity data and crack distribution data of the coal body and surrounding rock of the target fully-mechanized caving working face coal seam and its goaf, samples of coal body, roof rock, floor rock and other samples were drilled from the working face where the target tunnel was located in the coal mine underground and brought to the ground laboratory to be prepared into standard samples for mechanical testing to obtain the mechanical parameters of the rock numerical model, such as compressive strength, tensile strength, Poisson's ratio, etc. Specifically: long rock cylinders with a diameter of 50 mm were drilled from the underground working face to be mined, the tunnel roof, the tunnel floor and the coal pillar, respectively, and processed and prepared into cylindrical standard samples of different standard sizes such as Ф50×100mm and Ф50×250mm in the ground laboratory. After grouping the samples, the basic mechanical parameters of the standard samples were obtained by using the MTS triaxial loading tester, and the porosity and crack distribution results inside the standard samples were tested by using a mercury intrusion instrument and a rock CT machine. After the test, the compressive strength of the raw coal standard sample was 14.1MPa, the tensile strength was 1.0MPa, the Poisson's ratio was 0.31, and the elastic modulus was 1.1GPa; the porosity of the raw coal standard sample was 7.8%, the average crack opening was 90.1um, and the average crack volume was 2.1×10 7 um 3 , the crack degree is 0.83%.

[0052] In order to determine the technical parameters of hydraulic rock breaking, the top and bottom rock samples were drilled from the working face of the target tunnel in the coal mine and brought to the ground laboratory for water jet cutting experiments to determine the technical parameters of hydraulic rock breaking under different rock properties and different cutting depth requirements. Specifically: select a block sample with a cube shape and a volume of not less than 0.1m 3 In the laboratory, the block samples were processed into cubes with a side length of 20 cm, and experiments with different abrasive concentrations, jet pressures, nozzle radii and cutting times were carried out to obtain the optimal hydraulic rock breaking technical parameters.

[0053] Step 2, based on the mine data in Step 1, use Abaqus to construct a load-deformation finite element model of the adjacent working face goaf-coal pillar-tunnel-coal seam to be mined structure, and perform numerical simulation of the target working face mining process. Use mine data such as roof and top coal periodic fracture and collapse data, coal seam mining stress data, tunnel surrounding rock deformation and delamination during mine production to verify and optimize the load-deformation finite element model, and form a simulation model that meets the actual production conditions and monitoring data of the mine site. Based on the simulation model, perform numerical simulation of working face mining in the subsequent mining plan, calculate the tunnel area with high static load and large stress concentration coefficient and the roof fracture area of ​​the adjacent goaf, and identify the dangerous area of ​​the tunnel with the risk of impact mine pressure.

[0054] When conducting numerical simulation of the mining process of the target working face, firstly, the mine data including the geological conditions and mining conditions of the coal seam of the target fully-mechanized caving working face, as well as the basic mechanical parameter data of the coal body and surrounding rock of the coal seam and its goaf of the target fully-mechanized caving working face, the internal porosity data and the fracture distribution data are imported into the load-deformation finite element model; then, the vertical stress distribution of the structure is calculated according to the material mechanics theory to obtain the area with high elastic energy density of the roof; the grid of the area with high elastic energy density of the roof is encrypted; finally, the numerical simulation of the mining process of the target fully-mechanized caving working face is conducted to obtain the stress distribution, deformation characteristics and periodic collapse data of the roof and top coal of the target fully-mechanized caving working face and roadway.

[0055] Step 3: According to the coordinates of the dangerous area in the tunnel obtained in Step 2 and the coordinates of the far-field dynamic load source obtained by on-site microseismic monitoring, the hydraulic slot weakening anti-scouring conditions that can meet the anti-scouring requirements of the tunnel are set.

[0056] Hydraulic cutting weakening and anti-scouring must meet the following conditions:

[0057]

[0058] Where: Ω is the rock volume element; W is the energy transmitted from the dynamic load to the surrounding rock of the tunnel; U t is the strain energy of the rock element; ψ is the potential energy of the rock element; t is time; x j is the coordinate; σ j is the coordinate x j That is, the strain energy U of the rock element Ω t and body potential energy ψ for time t, coordinate x j and x j The stress σ j function.

[0059] in,

[0060] ψ=U s +U b +U cr

[0061] Where: U s is the energy consumed in the compaction stage; U b The energy consumed by the development of secondary cracks, the extension of primary cracks, and the final fracture surface energy; U cr It is the energy released in the form of heat energy, sound energy and electromagnetic radiation during the friction process of secondary crack fracture.

[0062] Volume element strain energy U t The energy W transmitted to the surrounding rock of the tunnel by the dynamic load satisfies the following conditions:

[0063]

[0064] Where: s is the static load of the tunnel surrounding rock; d is the dynamic load of the tunnel surrounding rock.

[0065] Initial static load of tunnel surrounding rock σ 0s With initial dynamic load σ 0d The following conditions must be met:

[0066]

[0067] Where: H n is the vertical height of coal strata at different burial depths; ρ n is the density of different rock layers; g is the acceleration of gravity; M is the mass of the collapsed rock layer; v is the collapse velocity of the rock layer; A is the impact area; λ is the Lame constant of the rock layer; w is the stress wave frequency of the impact load.

[0068] in,

[0069]

[0070] Where: c, k, G, u are the vertical damping coefficient, vertical spring coefficient, shear modulus, and Poisson's ratio of the rock layer, respectively; r is the equivalent radius of the collapsed rock mass; ρ is the density of the collapsed rock mass.

[0071] When there are hydraulic cutting joints in the tunnel surrounding rock, the static load σ s and the dynamic load σ of the tunnel surrounding rock d Different attenuation will occur, and the attenuated values ​​meet the following conditions:

[0072]

[0073] Where: r0 is the radius of the slit; r x is the distance between the surrounding rock position and the center of the cutting seam; C is the stress wave attenuation coefficient; x0 is the distance between the cutting seam and the dynamic load source; x is the distance between the surrounding rock and the dynamic load source.

[0074] For a certain volume element in the coal rock mass, if the energy W flowing into the volume element at a certain moment and the accumulated strain energy U t If the release rate of is greater than the absorption rate of its body energy ψ, the element becomes unstable and releases energy to the adjacent elements. If the adjacent elements cannot absorb this part of the energy, the adjacent elements become unstable and release energy, and the energy propagates along the dominant path of the medium to the weak surface of the excavation, eventually causing rock burst. Therefore, a multi-level hydraulic cutting simulation scheme can be planned according to the critical conditions of deformation instability of the surrounding rock of the tunnel and the rock mass of the far-field dynamic load source, and appropriate construction technical parameters for the far-field dynamic load source expansion cutting, the blocking cutting of the dynamic load propagation path, and the pressure relief cutting of the tunnel roof can be selected.

[0075] Step 4, based on the simulation model of Step 2 and the hydraulic slotting weakening anti-scouring conditions of Step 3, such as Figure 1 , Figure 2 As shown, firstly, simulated drilling is carried out for the dangerous area of ​​the tunnel in the simulation model, and hydraulic cutting is simulated to form a pressure relief slot, a blocking slot and a far-field dynamic load source expansion slot, wherein the pressure relief slot is located in the tunnel roof in the dangerous area of ​​the tunnel, the blocking slot is located on the line connecting the dangerous area of ​​the tunnel and the far-field dynamic load source, and the cutting surface of the blocking slot is perpendicular to the dynamic load propagation direction, the far-field dynamic load source expansion slot is in the same rock layer as the far-field dynamic load source in the vertical direction, and the far-field dynamic load source expansion slot is located near the far-field dynamic load source in the horizontal direction, then the dynamic load generated by the roof breaking is simulated at the far-field dynamic load source position, and the numerical simulation of the working face recovery of the subsequent mining plan is carried out at the same time, so as to obtain the optimal hydraulic cutting weakening and anti-collision technical parameters that can weaken the dynamic load and reduce the static load, and the hydraulic cutting weakening and anti-collision technical parameters include hydraulic cutting technical parameters such as hydraulic cutting orientation, distance, size, and number of layers.

[0076] The hydraulic cutting depth meets the following conditions:

[0077]

[0078] Where: p is the jet pressure; R is the nozzle radius; is the internal friction angle of the roof; τ is the shear strength of the roof rock; c is the cohesion of the roof rock.

[0079] When simulating hydraulic cutting to form a pressure relief groove, simulated drilling is carried out near the junction of the top coal and the coal pillar in the roadway, and simulated hydraulic cutting is carried out near the side of the roadway to reduce the vertical stress of the coal pillar roof on the roadway side. The downward extension range of the hydraulic cutting does not exceed the interface between the roof and the coal seam, and the upward extension range of the hydraulic cutting does not exceed the direct roof. After the roof is weakened, when the settlement height of the direct roof is ΔH, the roof pressure in the cutting area is

[0080]

[0081] Where: E2 is the direct top elastic modulus; L R1 is the radius of the pressure relief slot; θ1 is the elevation angle of the drill pipe of the pressure relief slot; H2 is the height of the direct roof; q is the pressure of the direct roof on the coal pillar.

[0082] Among them, the pressure relief slotted drill pipe elevation angle θ1 satisfies the following conditions:

[0083] θ1=arctan[(h1-h0) / l0]

[0084] In the formula: h1 is the tunnel height; h0 is the drilling rig height; l0 is the distance between the drilling rig and the coal pillar.

[0085] When obtaining the optimal hydraulic slot weakening and anti-shock technical parameters that can weaken dynamic load and reduce static load, by adjusting the hydraulic slot positions, slot spacing d, drill pipe elevation angle θ, hydraulic slot radius L R , the height dimension h0 of the drilling rig, the dimension l0 of the distance between the drilling rig and the goaf and other parameters were grouped for numerical simulation, and the dynamic load attenuation rate, the stress-strain change characteristics of the tunnel surrounding rock and the stress transfer characteristics of the coal pillar area under different hydraulic cutting parameters were obtained. The main control parameters of the multi-level hydraulic cutting weakening and anti-impact that affect the dynamic load attenuation rate, the stress-strain change characteristics of the tunnel surrounding rock and the stress transfer characteristics of the coal pillar area were determined. After verification and model optimization combined with the field observation results of the mine, an optimal scheme was formed that can effectively release the elastic energy of the rock mass near the dynamic load source, effectively block the dynamic load from propagating to the tunnel and release the static load of the tunnel surrounding rock.

[0086] Step 5, according to the optimal hydraulic cutting weakening and anti-bumping technical parameters obtained in Step 4, arrange drilling rigs in the target working face tunnel of the mine, drill holes in the tunnel surrounding rock roof, and use the high-pressure water jet ejected from the jet nozzle at the end of the drill pipe and sprayed along the radial direction of the drill pipe to impact and crush the roof rock. Through the rotation and retreat of the drill pipe, pressure relief slots, blocking slots and far-field dynamic load source broken expansion slots are constructed. By analogy, all pressure relief slots, blocking slots and far-field dynamic load source broken expansion slots are opened.

[0087] Step 6, based on the regional and real-time monitoring results of the tunnel surrounding rock in the target area monitored on site, determine the impact mine pressure hazard of the target tunnel surrounding rock before and after hydraulic cutting, evaluate and verify the anti-impact effect of hydraulic cutting, and repeat Step 4 for areas where the anti-impact effect does not meet the hydraulic cutting weakening anti-impact conditions of Step 3, adjust the optimal hydraulic cutting weakening anti-impact technical parameters obtained in Step 4, and perform the hydraulic cutting construction work of Step 5 again based on the adjusted and optimized optimal hydraulic cutting weakening anti-impact technical parameters until the anti-impact effect meets the hydraulic cutting weakening anti-impact conditions of Step 3.

[0088] When evaluating and verifying the anti-impact effect of hydraulic cutting: conduct microseismic monitoring in the tunnel after cutting, compare the frequency of the tunnel affected by mine earthquakes before and after cutting, compare and verify the measured data, theoretical model calculations and numerical simulation calculations, and evaluate the effect of far-field expansion cutting on reducing the far-field roof elastic energy density to inhibit the occurrence of mine earthquakes; compare the energy levels of mine earthquakes before and after cutting, compare and verify the measured data, theoretical model calculations and numerical simulation calculations, and evaluate the effect of blocking cutting on weakening dynamic loads; measure the deformation of the surrounding rock of the tunnel, compare the deformation of the surrounding rock in the cutting area and the uncut area, compare and verify the measured data, theoretical model calculations and numerical simulation calculations, and evaluate the effect of pressure-relief cutting on reducing the static load of the tunnel surrounding rock.

[0089] As the working face continues to advance, the mechanical properties and structure of the tunnel surrounding rock will change. On the basis of the multi-level hydraulic cutting weakening and anti-scouring operation that has been implemented, the areas where the anti-scouring effect does not meet the standard can be analyzed based on the real-time monitoring data of the tunnel surrounding rock deformation in the hydraulic cutting area and the microseismic monitoring data. The mechanical parameters of the tunnel surrounding rock and the hydraulic cutting technical parameters are corrected and grouped for numerical simulation, and the main control parameters that affect the multi-level hydraulic cutting weakening and anti-scouring effect under different geological conditions and different mining conditions are obtained again, and applied to the subsequent mining of the working face. That is, the real-time monitoring data continuously verifies and optimizes the numerical simulation model and the simulation model, and corrects the key parameters, and continues to perform the roof weakening operation of the multi-level hydraulic cutting to continuously maintain the anti-scouring ability of the tunnel at the end of the working face.

[0090] This method for weakening and preventing impact of the top plate at the end of the working face adopts the method of hydraulically cutting in the surrounding rock top plate in the dangerous area of ​​the tunnel at the end of the working face to form pressure relief grooves, blocking grooves and far-field dynamic load source fragmentation grooves to achieve the joint reduction of energy that can induce impact mining pressure disasters in three aspects: dynamic load generation conditions, dynamic load propagation path and surrounding rock static load, so as to improve the anti-impact capacity of the tunnel at the end of the working face: the pressure relief groove formed by hydraulic cutting in the tunnel top plate can release the static load of the tunnel surrounding rock, increase the degree of crushing of the hard tunnel top plate, reduce the strength of the top plate to transfer stress to deep rock, and at the same time increase the energy dissipation during the propagation of far-field dynamic loads; the blocking groove formed by hydraulic cutting on the dynamic load propagation path can reduce the strength of the top plate to transfer stress to deep rock and block the propagation of far-field dynamic loads to the tunnel; the far-field dynamic load source fragmentation groove formed by hydraulic cutting in the rock mass near the dynamic load source can reduce the elastic energy density of the rock mass near the dynamic load source, thereby reducing the possibility of roof rupture. This method of weakening the roof anti-impact at the end of the working face can effectively improve the anti-impact capacity of the working face end tunnel and significantly reduce the possibility of impact mine pressure disasters in the working face end tunnel. It is particularly suitable for solving the impact mine pressure problem in the working face end tunnel.

Claims

1. A method for weakening and preventing impact of the top plate at the end of a working surface, characterized in that: The anti-collision capability of the tunnel at the end of the working face is improved by hydraulically cutting the roof near the dynamic load source, the dynamic load transmission path and the tunnel roof, which specifically includes the following steps: Step 1: Obtain mine data, conduct on-site sampling, prepare standard samples for experiments, and obtain basic mechanical parameter data, internal porosity data, fracture distribution data, and optimal hydraulic rock breaking technical parameters of the target fully mechanized caving working face coal seam and its goaf. Step 2, based on the mine data in Step 1, construct a load-deformation finite element model of the adjacent working face goaf-coal pillar-tunnel-coal seam to be mined structure, and perform numerical simulation of the target working face mining process. Use the periodic fracture and collapse data of the roof and top coal during mine production, the coal seam mining stress data, and the tunnel surrounding rock deformation and separation data to verify and optimize the load-deformation finite element model, and form a simulation model that meets the actual production conditions and monitoring data of the mine site. Based on the simulation model, perform numerical simulation of the working face mining of the subsequent mining plan, calculate the tunnel area with high static load and large stress concentration coefficient and the roof fracture area of ​​the adjacent goaf, and identify the dangerous area of ​​the tunnel with the danger of impact mine pressure; Step 3: According to the coordinates of the dangerous area of ​​the tunnel obtained in Step 2 and the coordinates of the far-field dynamic load source obtained by on-site microseismic monitoring, the hydraulic slot weakening anti-scouring conditions that can meet the anti-scouring requirements of the tunnel are set; Hydraulic cutting weakening and anti-scouring must meet the following conditions: Where: Ω is the rock volume element; W is the energy transmitted from the dynamic load to the surrounding rock of the tunnel; U t is the strain energy of the rock element; ψ is the potential energy of the rock element; t is time; x j is the coordinate; σ j is the coordinate x j Stress at location; in, ψ=U s +U b +U cr Where: U s is the energy consumed in the compaction stage; U b The energy consumed by the development of secondary cracks, the extension of primary cracks, and the final fracture surface energy; U cr It is the energy released during the friction process of secondary crack fracture; Volume element strain energy U t The energy W transmitted to the surrounding rock of the tunnel by the dynamic load satisfies the following conditions: Where: s is the static load of the tunnel surrounding rock; d is the dynamic load of the tunnel surrounding rock; Initial static load of tunnel surrounding rock σ 0s With initial dynamic load σ 0d The following conditions must be met: Where: H n is the vertical height of coal strata at different burial depths; ρ n is the density of different rock layers; g is the acceleration of gravity; M is the mass of the collapsed rock layer; v is the collapse velocity of the rock layer; A is the impact area; λ is the Lame constant of the rock layer; w is the stress wave frequency of the impact load; in, Where: c, k, G, u are the vertical damping coefficient, vertical spring coefficient, shear modulus, and Poisson's ratio of the rock layer respectively; r is the equivalent radius of the collapsed rock mass; ρ is the density of the collapsed rock mass; When there are hydraulic cutting joints in the tunnel surrounding rock, the static load σ s and the dynamic load σ of the tunnel surrounding rock d Different attenuation will occur, and the attenuated values ​​meet the following conditions: Where: r0 is the radius of the slit; r x is the distance between the surrounding rock position and the center of the cutting seam; C is the stress wave attenuation coefficient; x0 is the distance between the cutting seam and the dynamic load source; x is the distance between the surrounding rock and the dynamic load source; Step 4, based on the simulation model of Step 2 and the hydraulic cutting weakening anti-scouring conditions of Step 3, firstly, simulate drilling and simulate hydraulic cutting for the dangerous area of ​​the tunnel in the simulation model to form pressure relief slots, blocking slots and far-field dynamic load source broken expansion slots, wherein the pressure relief slots are located in the tunnel roof in the dangerous area of ​​the tunnel, the blocking slots are located on the line connecting the dangerous area of ​​the tunnel and the far-field dynamic load source, and the cutting surface of the blocking slots is perpendicular to the dynamic load propagation direction, the far-field dynamic load source broken expansion slots are in the same rock layer as the far-field dynamic load source in the vertical direction, and the far-field dynamic load source broken expansion slots are located near the far-field dynamic load source in the horizontal direction, then simulate the dynamic load generated by the roof breaking at the far-field dynamic load source position, and simultaneously carry out numerical simulation of the working face recovery of the subsequent mining plan, so as to obtain the optimal hydraulic cutting weakening anti-scouring technical parameters that can weaken the dynamic load and reduce the static load; Step 5, according to the optimal hydraulic cutting weakening and anti-bumping technical parameters obtained in Step 4, a drilling rig is arranged in the target working face tunnel of the mine, and holes are drilled into the tunnel surrounding rock roof respectively, and the top rock is crushed by impacting the high-pressure water jet ejected from the jet nozzle at the end of the drill pipe and ejected along the radial direction of the drill pipe. Through the rotation and retreat of the drill pipe, pressure relief slots, blocking slots and far-field dynamic load source broken expansion slots are constructed, and so on, all pressure relief slots, blocking slots and far-field dynamic load source broken expansion slots are completed; Step 6, based on the regional and real-time monitoring results of the tunnel surrounding rock in the target area monitored on site, determine the impact mine pressure hazard of the target tunnel surrounding rock before and after hydraulic cutting, evaluate and verify the anti-impact effect of hydraulic cutting, and repeat Step 4 for areas where the anti-impact effect does not meet the hydraulic cutting weakening anti-impact conditions of Step 3, adjust the optimal hydraulic cutting weakening anti-impact technical parameters obtained in Step 4, and perform the hydraulic cutting construction work of Step 5 again based on the adjusted and optimized optimal hydraulic cutting weakening anti-impact technical parameters until the anti-impact effect meets the hydraulic cutting weakening anti-impact conditions of Step 3.

2. The method for weakening and preventing impact of the top plate at the end of the working face according to claim 1 is characterized in that: When numerically simulating the mining process of the target working face in Step 2, firstly, the mine data including the geological conditions and mining conditions of the coal seam of the target fully-mechanized caving working face, as well as the basic mechanical parameter data of the coal body and surrounding rock of the coal seam and its goaf of the target fully-mechanized caving working face, the internal porosity data and the fracture distribution data are imported into the load-deformation finite element model; then, the vertical stress distribution is calculated to obtain the area with high elastic energy density of the roof; the mesh of the area with high elastic energy density of the roof is encrypted; finally, the numerical simulation of the mining process of the target fully-mechanized caving working face is carried out to obtain the stress distribution, deformation characteristics and periodic collapse data of the roof and top coal of the target fully-mechanized caving working face and roadway.

3. The method for weakening and preventing impact of the top plate of the working face end according to claim 1, characterized in that: When simulating hydraulic cutting in dangerous areas of tunnels in the simulation model in Step 4, the hydraulic cutting depth meets the following conditions: Where: p is the jet pressure; R is the nozzle radius; is the internal friction angle of the roof; τ is the shear strength of the roof rock; c is the cohesion of the roof rock.

4. The method for weakening and preventing impact of the top plate at the end of the working face according to claim 1, characterized in that: When simulating hydraulic cutting to form a pressure relief groove in Step 4, simulated drilling is carried out near the junction of the top coal and the coal pillar in the tunnel, and simulated hydraulic cutting is carried out near the side of the tunnel. The downward extension range of the hydraulic cutting does not exceed the interface between the roof and the coal seam, and the upward extension range of the hydraulic cutting does not exceed the direct roof.

5. The method for weakening and preventing impact of the top plate at the end of the working face according to claim 1, characterized in that: When simulating hydraulic cutting to form a pressure relief slot in Step 4, after the roof is weakened, when the direct top settlement height is ΔH, the roof pressure in the cutting area is Where: E2 is the direct top elastic modulus; L R1 is the radius of the pressure relief slot; θ1 is the elevation angle of the pressure relief slot drill pipe; H2 is the height of the direct roof; q is the pressure of the direct roof on the coal pillar; Among them, the pressure relief slotted drill pipe elevation angle θ1 satisfies the following conditions: θ1=arctan[(h1-h0) / l0] In the formula: h1 is the tunnel height; h0 is the drilling rig height; l0 is the distance between the drilling rig and the coal pillar.

6. The method for weakening and preventing impact of the top plate of the working face end according to claim 1, characterized in that: When obtaining the optimal hydraulic slotting anti-collision technical parameters that can weaken dynamic loads and reduce static loads in Step 4, the dynamic load attenuation rate, the stress-strain change characteristics of the tunnel surrounding rock and the stress transfer characteristics of the coal pillar area under different hydraulic slotting parameters are obtained by grouping numerical simulation of different hydraulic slotting parameters. The main control parameters of hydraulic slotting anti-collision that affect the dynamic load attenuation rate, the stress-strain change characteristics of the tunnel surrounding rock and the stress transfer characteristics of the coal pillar area are judged and obtained. After verification and model optimization combined with the field observation results of the mine, the optimal hydraulic slotting anti-collision technical parameters are formed.

7. The method for weakening and preventing impact of the top plate of the working face end according to claim 1, characterized in that: When evaluating and verifying the anti-blowout effect of hydraulic cutting in Step 6, microseismic monitoring is carried out in the tunnel after cutting, the frequency of the tunnel affected by mine earthquakes before and after cutting is compared, the energy level of mine earthquakes before and after cutting is compared, the deformation of the surrounding rock of the tunnel is measured, and the deformation of the surrounding rock in the cutting area and the uncut area is compared. The measured data, theoretical model calculations and numerical simulation calculations are compared and verified, and the effect of far-field expansion cutting on reducing the far-field roof elastic energy density to suppress mine earthquakes is evaluated, the effect of blocking cutting on weakening dynamic loads is evaluated, and the effect of pressure-relief cutting on reducing the static load of the tunnel surrounding rock is evaluated.

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