A method for weakening and preventing impact on the top plate of the working face

Abstract

This invention discloses a method for weakening and preventing rockburst at the end of a working face. It improves the rockburst resistance of the working face end roadway by hydraulically cutting grooves in the roof of the surrounding rock in the hazardous area of ​​the roadway. The pressure-relieving grooves formed by hydraulic cutting release the static load of the surrounding rock. The blocking grooves formed by hydraulic cutting along the dynamic load propagation path block the propagation of the dynamic load to the roadway. The far-field dynamic load source fracture grooves formed by hydraulic cutting in the rock mass near the dynamic load source reduce the elastic energy density of the rock mass near the dynamic load source and decrease the possibility of roof fracture. This invention can simultaneously reduce the energy that can induce rockburst disasters from three aspects: the generation conditions of dynamic load, the propagation path of dynamic load, and the static load of the surrounding rock. It is particularly suitable for solving the rockburst problem in the working face end roadway.

Description

Technical Field

[0001] This invention relates to a method for weakening and preventing rockburst at the end of a working face. Specifically, it is a method for weakening and preventing rockburst at the end of a working face by reducing the risk of rockburst in roadways in three aspects: the generation conditions of dynamic load, the propagation path of dynamic load, and the static load of surrounding rock. It belongs to the field of coal mine safety mining technology. Background Technology

[0002] Coal mining in my country is primarily underground. With the increasing depletion of shallow resources, coal mining is gradually extending to deeper underground areas. This leads to increased in-situ stress in coal seams, enhanced mining disturbance, and aggravated rockburst manifestations in coal mine production. Extensive research, both domestically and internationally, has been conducted on the manifestation patterns and disaster mechanisms of rockburst in hard roofs. Rockburst is an instability phenomenon induced when the energy released by the coal-rock system exceeds the energy consumed. Rockburst will not occur when the stress on the coal-rock mass does not exceed its ultimate strength, nor when the rate of elastic energy release in the coal-rock system is less than the rate of energy dissipation during crack propagation. However, mining activities cause a readjustment of stress and energy in the surrounding rock of the mining space. When the static load on the surrounding rock and the dynamic load caused by external dynamic loads exceed the critical load for coal-rock rockburst failure, the coal-rock system becomes unstable and fails, thus inducing rockburst. Especially when the roof of the working face is hard, as the working face advances, the roof above the goaf becomes too long, and a large amount of elastic energy tends to accumulate in the roof at the end of the working face. If a sudden fracture occurs under the action of mining stress or impact load generated by geological activity, it may induce a roof accident in the roadway and then induce rockburst.

[0003] Existing technologies for preventing rockbursts mainly fall into four categories: optimizing roadway layout, strengthening support, backfilling for rockburst prevention, and weakening the hard roof. Optimizing roadway layout effectively utilizes the stress-reducing zone of the goaf to reduce the static load on the roadway. Strengthening support, using high-prestressed rock bolts, can effectively control the deformation of the surrounding rock. However, neither of these methods fundamentally solves the high-stress zone in the roof caused by the superposition of dynamic and static loads. While backfilling for rockburst prevention can effectively solve the problem of suspended roofs, it is costly. Therefore, current measures for preventing rockbursts primarily focus on weakening the strength of the hard roof. Traditional roof weakening methods include water injection to weaken the roof, borehole decompression, hydraulic fracturing, and borehole blasting pre-fracturing. Water injection to weaken the roof has significant limitations and is not effective for rock masses with low porosity, low water absorption, and incomplete fracture development. Drilling can reduce the stress concentration of the surrounding coal and rock mass, but the influence 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-fracturing has a large stress relief range, but the operation is difficult and the dynamic load generated may induce shock pressure. Due to safety and environmental protection restrictions, it is difficult to apply it on a large scale. Summary of the Invention

[0004] To address the problems existing in the prior art, this invention provides a method for weakening and preventing rockburst at the working face end. This method can reduce the energy that can induce rockburst disasters by considering three aspects: the generation conditions of dynamic loads, the propagation path of dynamic loads, and the static load of the surrounding rock. As a result, the load borne by the working face end roadway is lower than the critical value that can induce rockburst disasters, and the rockburst prevention capability of the working face end roadway is improved. This method is particularly suitable for solving the rockburst problem in the working face end roadway.

[0005] To achieve the above objectives, this method for weakening and preventing scour at the working face end roof improves the scour resistance of the roadway at the working face end by hydraulically cutting grooves in the roof near the dynamic load source, the dynamic load propagation path, and the roadway roof. Specifically, it includes the following steps:

[0006] Step 1: Obtain mine data and 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 technology parameters of the coal seam and surrounding rock of the target fully mechanized longwall face and its goaf.

[0007] Step 2: Based on the mine data from Step 1, construct a finite element model of the load-deformation structure of the adjacent working face goaf-coal pillar-roadway-coal seam to be mined in this working face, and conduct numerical simulation of the target working face mining process. Utilize data on periodic fracture and collapse of the roof and top coal during mine production, coal seam mining stress data, and roadway surrounding rock deformation and delamination data to verify and optimize the parameters of this load-deformation finite element model, forming a simulation model that conforms to the actual production conditions and monitoring data of the mine site. Based on the simulation model, conduct numerical simulation of the working face mining in the subsequent mining plan, calculate the roadway areas with high static load and large stress concentration coefficient, and the roof fracture areas of adjacent goaf areas, and identify the dangerous roadway areas with the risk of rockburst.

[0008] Step 3: Based on the coordinates of the dangerous area of ​​the roadway obtained in Step 2 and the coordinates of the far-field dynamic load source obtained from on-site microseismic monitoring, set hydraulic slotting weakening anti-scour conditions that can meet the anti-scour requirements of the roadway.

[0009] Hydraulic kerving weakening and erosion protection must meet the following conditions:

[0010]

[0011] In the formula: Ω represents the rock mass element; W represents the energy of dynamic load propagating to the surrounding rock of the roadway; U t denoted as ψ, representing the strain energy of the rock volume element; t represents time; x represents the volume potential energy of the rock volume element. j For coordinates; σ j Let x be the coordinate j Stress at a location;

[0012] in,

[0013] ψ=U s +U b +U cr

[0014] In the formula: U s Energy consumed during the compaction stage; U b The energy consumed by the development of secondary cracks, the propagation of primary cracks, and the final fracture to form surface energy; U cr This refers to the energy released during the secondary crack fracture friction process;

[0015] Volumetric strain energy U t The energy W transmitted from the dynamic load to the surrounding rock of the roadway satisfies the following condition:

[0016]

[0017] In the formula: σ s σ is the static load on the surrounding rock of the tunnel; d The dynamic load on the surrounding rock of the tunnel;

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

[0019]

[0020] Where: H n ρ represents the vertical height of coal and rock strata at different burial depths. n denoted as density of different rock layers; g is the acceleration due to 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 Lamé constant of the rock layer; w is the stress wave frequency of the impact load.

[0021] in,

[0022]

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

[0024] When hydraulic cuts exist in the surrounding rock of the tunnel, the static load σ of the surrounding rock is... s Dynamic load σ of the surrounding rock of the tunnel d All will experience different degrees of decay, and the decayed values ​​will satisfy the following conditions:

[0025]

[0026] In the formula: r0 is the radius of the kerf; r x denoted as , where is the distance between the surrounding rock location and the center of the cut; C is the stress wave attenuation coefficient; x0 is the distance between the cut 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 in Step 2 and the hydraulic slotting weakening and anti-scour conditions in Step 3, firstly, simulated drilling is performed in the dangerous area of ​​the roadway in the simulation model, and simulated hydraulic slotting to form pressure relief slots, blocking slots, and far-field dynamic load source fracture expansion slots. Among them, the pressure relief slots are located in the roof of the roadway in the dangerous area, the blocking slots are located on the line connecting the dangerous area of ​​the roadway and the far-field dynamic load source, and the slotting surface of the blocking slots is perpendicular to the direction of dynamic load propagation. The far-field dynamic load source fracture expansion slots are in the same rock layer as the far-field dynamic load source in the vertical direction, and are located near the far-field dynamic load source in the horizontal direction. Then, the dynamic load generated by roof fracture is simulated at the location of the far-field dynamic load source, and the numerical simulation of the working face mining of the subsequent mining plan is carried out to obtain the optimal hydraulic slotting weakening and anti-scour technology parameters that can weaken the dynamic load and reduce the static load.

[0028] Step 5: Based on the optimized hydraulic slotting weakening and anti-scour technology parameters obtained in Step 4, drilling rigs are deployed in the roadway of the target working face of the mine to drill holes into the roof of the roadway surrounding rock. High-pressure water jets, sprayed from the nozzles at the end of the drill rods and radially along the drill rods, are used to impact and break the roof rock. Through the rotation and retraction of the drill rods, pressure relief slots, blocking slots, and far-field dynamic source fracture expansion slots are formed. This process is repeated to complete the opening of all pressure relief slots, blocking slots, and far-field dynamic source fracture expansion slots.

[0029] Step 6: Based on the regional and real-time monitoring results of the surrounding rock of the target roadway in the field monitoring area, determine the risk of rockburst in the target roadway before and after hydraulic cutting, evaluate and verify the anti-scour effect of hydraulic cutting, and for areas where the anti-scour effect does not meet the conditions for weakening anti-scour in Step 3, repeat Step 4 and adjust the optimized hydraulic cutting weakening anti-scour technical parameters obtained in Step 4. Based on the adjusted and optimized optimized hydraulic cutting weakening anti-scour technical parameters, perform the hydraulic cutting construction operation in Step 5 again until the anti-scour effect meets the conditions for weakening anti-scour in Step 3.

[0030] Furthermore, in Step 2, when performing numerical simulation of the target longwall face mining process, the following steps are taken: First, mine data including the geological conditions and mining conditions of the coal seam in the target longwall face, as well as the basic mechanical parameters, internal porosity data, and fracture distribution data of the coal body and surrounding rock of the target longwall face and its goaf are imported into the load-deformation finite element model. Then, the vertical stress distribution is calculated to obtain the areas with high roof elastic energy density. The mesh is then refined in the areas with high roof elastic energy density. Finally, numerical simulation of the target longwall face mining process is performed to obtain the stress distribution, deformation characteristics, and periodic collapse data of the roof and top coal in the target longwall face and roadway.

[0031] Furthermore, in Step 4, when simulating hydraulic cutting in the dangerous area of ​​the roadway in the simulation model, the hydraulic cutting depth must meet the following condition:

[0032]

[0033] In the formula: p is the jet pressure; R is the nozzle radius; τ 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, in Step 4, when simulating the formation of pressure relief grooves by hydraulic cutting, simulated drilling is carried out near the junction of the roof coal and the coal pillar in the roadway, and simulated hydraulic cutting is carried out on the side close to the roadway. The downward extension range of the hydraulic cutting does not exceed the junction of the roof and the coal seam, and the upward extension range of the hydraulic cutting does not exceed the direct roof.

[0035] Furthermore, in Step 4, when simulating the formation of a pressure relief groove through hydraulic cutting, after the top plate is weakened, when the direct roof settlement height is ΔH, the roof pressure in the cutting area is...

[0036]

[0037] In the formula: E2 is the direct top elastic modulus; L R1 θ1 is the radius of the pressure relief cut; H2 is the height of the immediate roof; q is the pressure of the immediate roof on the coal pillar.

[0038] The elevation angle θ1 of the pressure relief slotted drill pipe satisfies the following condition:

[0039] θ1 = arctan[(h1 - h0) / l0]

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

[0041] Furthermore, in Step 4, when obtaining the optimal hydraulic slotting weakening and anti-scour technology parameters that can reduce dynamic load and decrease static load, group numerical simulations are performed on different parameters of the hydraulic slotting to obtain the dynamic load attenuation rate, roadway surrounding rock stress-strain change characteristics, and coal pillar area stress transfer characteristics under different hydraulic slotting parameter conditions. The main control parameters of hydraulic slotting anti-scour that affect the dynamic load attenuation rate, roadway surrounding rock stress-strain change characteristics, and coal pillar area stress transfer characteristics are determined and obtained. After verification and model optimization based on the mine field observation results, the optimal hydraulic slotting weakening and anti-scour technology parameters are formed.

[0042] Furthermore, in Step 6, when evaluating and verifying the anti-scouring effect of hydraulic slotting, microseismic monitoring is conducted in the roadway after slotting. The frequency of mine-induced tremors affecting the roadway before and after slotting is compared, the energy level of mine-induced tremors before and after slotting is compared, the deformation of the surrounding rock in the roadway is measured, and the deformation of the surrounding rock in the slotted area and the unslotted area is compared. The measured data, theoretical model calculations and numerical simulation calculations are compared and verified to evaluate the effect of far-field dilatation slotting in reducing the elastic energy density of the far-field roof and suppressing the occurrence of mine-induced tremors, the effect of stalling slotting in weakening dynamic loads, and the effect of pressure relief slotting in reducing the static load of the roadway surrounding rock.

[0043] Compared with existing technologies, this method for weakening and preventing scour at the working face end of the roadway improves the scour resistance of the roadway by hydraulically cutting grooves in the roof of the surrounding rock in the dangerous area of ​​the roadway at the working face end. This is achieved by creating pressure relief grooves, blocking grooves, and far-field dynamic load source fracture expansion grooves. The pressure relief grooves created by hydraulic cutting in the roadway roof can release the static load of the roadway surrounding rock, increase the degree of fragmentation of the hard roof, reduce the roof strength and transfer stress to the deeper rock, and at the same time increase the energy dissipation during the propagation of far-field dynamic loads. The blocking grooves created by hydraulic cutting in the path of dynamic load propagation can reduce the roof strength, transfer stress to the deeper rock, and block the propagation of far-field dynamic loads to the roadway. The far-field dynamic load source fracture expansion grooves created 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 fracture. It can reduce the energy that can induce rockburst disasters by considering the dynamic load generation conditions, the dynamic load propagation path, and the static load of the surrounding rock. It can reduce the load on the working face end roadway below the critical value that can induce rockburst disasters, improve the rockburst resistance of the working face end roadway, and thus significantly reduce the possibility of rockburst disasters occurring in the working face end roadway. At the same time, the hydraulic cutting construction process is relatively refined and the amount of engineering work 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 rockburst problem in the working face end roadway. Attached Figure Description

[0044] Figure 1 This is a top view of the hydraulic cavity-making construction method of the present invention;

[0045] Figure 2 A cross-sectional view along the working face. Detailed Implementation

[0046] The method for weakening and preventing rockburst at the end of this working face reduces the energy that can induce rockburst disasters 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 and reduce the possibility of roof fracture; hydraulic cutting is performed on the dynamic load propagation path to block the propagation of far-field dynamic loads to the roadway; and hydraulic cutting is performed on the roof of the coal pillar to release the static load of the roadway surrounding rock.

[0047] The invention will be further explained below using an isolated working face roadway in a coal mine as an example, in conjunction with the accompanying drawings.

[0048] The method for weakening and preventing impact on the top plate of the working face includes the following steps:

[0049] Step 1: Obtain mine data and 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 technology parameters of the coal seam and surrounding rock of the target fully mechanized longwall face and its goaf.

[0050] The average burial depth of the integrated top-coal caving longwall face in this coal mine is 450m, the average coal seam dip angle is 6°, the average coal seam thickness is 7m, the immediate roof thickness is 15.5m, and it is separated from the goaf by a 15m wide coal pillar. The return airway of the longwall face 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 methods for the return airway are full anchor cable support and single hydraulic prop support. According to the monitoring data analysis, the length of the goaf roof overhang during the target longwall face mining period is between 15 and 35m, and the periodic pressure step distance of the longwall face is 20m.

[0051] To accurately obtain basic mechanical parameters, internal porosity data, and fracture distribution data of the coal seam and surrounding rock in the target fully mechanized longwall face and its goaf, samples of coal body, roof rock, and floor rock were drilled from the working face where the target roadway is located underground in the coal mine. These samples were then transported to a surface laboratory and prepared into standard samples for mechanical testing to obtain the mechanical parameters of the rock numerical model, such as compressive strength, tensile strength, and Poisson's ratio. Specifically, long rock cylinders with a diameter of 50 mm were drilled from the working face to be mined, the roadway roof, the roadway floor, and the coal pillar. These were then processed in the surface laboratory into standard cylindrical samples of different standard sizes, such as Ф50×100 mm and Ф50×250 mm. After grouping the samples, the basic mechanical parameters of the standard samples were obtained using an MTS triaxial loading tester. The porosity and fracture distribution inside the standard samples were tested using a mercury porosimeter and a rock CT scanner, respectively. After testing, the compressive strength of the raw coal standard sample was 14.1 MPa, the tensile strength was 1.0 MPa, the Poisson's ratio was 0.31, and the elastic modulus was 1.1 GPa. The porosity of the raw coal standard sample was 7.8%, the average fracture aperture was 90.1 μm, and the average fracture volume was 2.1 × 10⁻⁶. 7 um 3 The porosity is 0.83%.

[0052] To determine the technical parameters for hydraulic rock breaking, roof and floor rock samples were drilled from the working face of the target roadway in the coal mine and transported to a surface laboratory for water jet slotting experiments. The experiments aimed to determine the technical parameters for hydraulic rock breaking under different lithologies and slotting depth requirements. Specifically, approximately cubic block samples with a volume of not less than 0.1 m³ were selected underground. 3 In the laboratory, the block sample was processed into a cube with a side length of 20cm, and experiments were conducted with different abrasive concentrations, jet pressures, nozzle radii and cutting times to obtain the optimal hydraulic rock breaking technology parameters.

[0053] Step 2: Based on the mine data from Step 1, a finite element model of the load-deformation structure of the adjacent working face goaf-coal pillar-roadway-coal seam to be mined in this working face is constructed using Abaqus. Numerical simulation of the target working face mining process is then performed. The load-deformation finite element model is verified and its parameters are optimized using mine data such as roof and top coal periodic fracture and collapse data, coal seam mining stress data, and roadway surrounding rock deformation and delamination data. This forms a simulation model that conforms to the actual production conditions and monitoring data of the mine site. Based on the simulation model, numerical simulation of the working face mining in the subsequent mining plan is performed. The roadway areas with high static load and high stress concentration coefficient and the roof fracture areas of adjacent goaf areas are calculated, and the dangerous roadway areas with the risk of rockburst are identified.

[0054] When conducting numerical simulation of the target longwall face mining process, the following steps are taken: First, mine data including the geological conditions and mining conditions of the coal seam in the target longwall face, as well as the basic mechanical parameters, internal porosity data, and fracture distribution data of the coal body and surrounding rock in the target longwall face and its goaf are imported into the load-deformation finite element model. Then, the vertical stress distribution of the structure is calculated according to the theory of mechanics of materials to obtain the areas with high roof elastic energy density. The mesh is then refined in the areas with high roof elastic energy density. Finally, numerical simulation of the target longwall face mining process is conducted to obtain the stress distribution, deformation characteristics, and periodic collapse data of the roof and top coal in the target longwall face and roadway.

[0055] Step 3: Based on the coordinates of the dangerous area of ​​the roadway obtained in Step 2 and the coordinates of the far-field dynamic load source obtained from on-site microseismic monitoring, set hydraulic slotting weakening anti-scour conditions that can meet the anti-scour requirements of the roadway.

[0056] Hydraulic kerving weakening and erosion protection must meet the following conditions:

[0057]

[0058] In the formula: Ω represents the rock mass element; W represents the energy of dynamic load propagating to the surrounding rock of the roadway; U t denoted as ψ, representing the strain energy of the rock volume element; t represents time; x represents the volume potential energy of the rock volume element. j For coordinates; σ j Let x be the coordinate j The stress at a location. That is, the strain energy U of the rock mass element Ω. t The volume potential energy ψ is given by time t and coordinate x. j and x j Stress σ at the point j The function.

[0059] in,

[0060] ψ=U s +U b +U cr

[0061] In the formula: U s Energy consumed during the compaction stage; U b The energy consumed by the development of secondary cracks, the propagation of primary cracks, and the final fracture to form surface energy; U cr This refers to the energy released during the secondary crack fracture friction process in the form of heat, sound, and electromagnetic radiation.

[0062] Volumetric strain energy U t The energy W transmitted from the dynamic load to the surrounding rock of the roadway satisfies the following condition:

[0063]

[0064] In the formula: σ s σ is the static load on the surrounding rock of the tunnel; d This refers to the dynamic load on the surrounding rock of the tunnel.

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

[0066]

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

[0068] in,

[0069]

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

[0071] When hydraulic cuts exist in the surrounding rock of the tunnel, the static load σ of the surrounding rock is... s Dynamic load σ of the surrounding rock of the tunnel d All will experience different degrees of decay, and the decayed values ​​will satisfy the following conditions:

[0072]

[0073] In the formula: r0 is the radius of the kerf; r x denoted as , where is the distance between the surrounding rock location and the center of the cut; C is the stress wave attenuation coefficient; x0 is the distance between the cut and the dynamic load source; and x is the distance between the surrounding rock and the dynamic load source.

[0074] For a specific volume element within a coal and rock mass, if at a certain moment the energy W flowing into that volume element and the accumulated strain energy U... t If the release rate of energy exceeds the absorption rate of its volumetric energy ψ, the volume element becomes unstable and releases energy to neighboring volume elements. If neighboring volume elements cannot absorb this energy, they become unstable and release energy, which propagates along the dominant path of the medium to the weakest surface of the excavation, ultimately causing rockburst. Therefore, a multi-level hydraulic slotting simulation scheme can be planned based on the deformation and instability critical conditions of the surrounding rock of the tunnel and the far-field dynamic load source rock mass, and appropriate construction technical parameters for far-field dynamic load source dilatation slotting, dynamic load propagation path blocking slotting, and tunnel roof pressure relief slotting can be selected.

[0075] Step 4, based on the simulation model of Step 2 and the hydraulic slit weakening anti-scour conditions of Step 3, such as... Figure 1 , Figure 2 As shown, firstly, simulated drilling is performed on the dangerous area of ​​the roadway in the simulation model, and simulated hydraulic cutting to form pressure relief slots, blocking slots, and far-field dynamic load source fracture expansion slots. The pressure relief slots are located in the roof of the dangerous area of ​​the roadway, the blocking slots are located on the line connecting the dangerous area of ​​the roadway and the far-field dynamic load source, and the cutting surface of the blocking slots is perpendicular to the direction of dynamic load propagation. The far-field dynamic load source fracture expansion slots are in the same rock layer as the far-field dynamic load source in the vertical direction, and are located near the far-field dynamic load source in the horizontal direction. Then, the dynamic load generated by roof fracture is simulated at the location of the far-field dynamic load source, and the numerical simulation of the working face mining of the subsequent mining plan is carried out to obtain the optimal hydraulic cutting weakening and anti-scour technology parameters that can weaken the dynamic load and reduce the static load. The hydraulic cutting weakening and anti-scour technology parameters include hydraulic cutting orientation, distance, size, number of layers, and other hydraulic cutting technology parameters.

[0076] The hydraulic cut depth must meet the following conditions:

[0077]

[0078] In the formula: p is the jet pressure; R is the nozzle radius; τ 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 the formation of a pressure-relieving slot using hydraulic cutting, simulated drilling is performed near the interface between the roof coal and the coal pillar in the roadway, and simulated hydraulic cutting is performed on the side close to the roadway to reduce the vertical stress on the roof of the coal pillar on the roadway side. The downward extension of the hydraulic cutting does not exceed the interface between the roof and the coal seam, and the upward extension of the hydraulic cutting does not exceed the immediate roof. After the roof is weakened, when the immediate roof settlement height is ΔH, the roof pressure in the cutting area is...

[0080]

[0081] In the formula: E2 is the direct top elastic modulus; L R1 θ1 is the radius of the pressure relief cut; H2 is the height of the direct roof; q is the pressure of the direct roof on the coal pillar.

[0082] The elevation angle θ1 of the pressure relief slotted drill pipe satisfies the following condition:

[0083] θ1 = arctan[(h1 - h0) / l0]

[0084] In the formula: h1 is the roadway 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 slit weakening and anti-scour technology parameters that can reduce dynamic load and decrease static load, different hydraulic slit positions, slit spacing d, drill pipe elevation angle θ, and hydraulic slit radius L are considered. R Numerical simulations were performed in groups using parameters such as the height h0 of the drilling rig and the distance l0 between the drilling rig and the goaf. This yielded dynamic load attenuation rate, stress-strain variation characteristics of the surrounding rock in the roadway, and stress transfer characteristics of the coal pillar area under different hydraulic cutting parameters. The main control parameters for multi-level hydraulic cutting weakening and scour prevention that affect the dynamic load attenuation rate, stress-strain variation characteristics of the surrounding rock in the roadway, and stress transfer characteristics of the coal pillar area were determined. After verification and model optimization based on mine field observation results, an optimal scheme was formed that can effectively release the elastic energy of the rock mass near the dynamic load source, effectively block the propagation of dynamic load to the roadway, and release the static load of the surrounding rock in the roadway.

[0086] Step 5: Based on the optimized hydraulic slotting weakening and anti-scour technology parameters obtained in Step 4, drilling rigs are deployed in the target working face roadway of the mine to drill holes into the roof of the roadway surrounding rock. High-pressure water jets, sprayed radially along the drill rod through the jet nozzle at the end of the drill rod, are used to impact and break the roof rock. Through the rotation and retraction of the drill rod, pressure relief slots, blocking slots, and far-field dynamic source fracture expansion slots are formed. This process is repeated to complete the opening of all pressure relief slots, blocking slots, and far-field dynamic source fracture expansion slots.

[0087] Step 6: Based on the regional and real-time monitoring results of the surrounding rock of the target roadway in the field monitoring area, determine the risk of rockburst in the target roadway before and after hydraulic cutting, evaluate and verify the anti-scour effect of hydraulic cutting, and for areas where the anti-scour effect does not meet the conditions for weakening anti-scour in Step 3, repeat Step 4 and adjust the optimized hydraulic cutting weakening anti-scour technical parameters obtained in Step 4. Based on the adjusted and optimized optimized hydraulic cutting weakening anti-scour technical parameters, perform the hydraulic cutting construction operation in Step 5 again until the anti-scour effect meets the conditions for weakening anti-scour in Step 3.

[0088] When evaluating and verifying the anti-scouring effect of hydraulic slotting: microseismic monitoring is conducted in the roadway after slotting, and the frequency of mine-induced tremors affecting the roadway before and after slotting is compared. The measured data, theoretical model calculations, and numerical simulations are compared and verified to evaluate the effect of far-field dilatation slotting in reducing the elastic energy density of the far-field roof and suppressing the occurrence of mine-induced tremors; the energy levels of mine-induced tremors before and after slotting are compared and verified to evaluate the effect of slotting in weakening dynamic loads; the deformation of the surrounding rock in the roadway is measured, and the deformation of the surrounding rock in the slotted area and the unslotted area is compared and verified to evaluate the effect of pressure-relief slotting in reducing the static load of the surrounding rock in the roadway.

[0089] As the working face continues to advance, the mechanical properties and structure of the surrounding rock in the roadway will change. Based on the multi-level hydraulic slotting weakening and anti-scour operations already implemented, areas where the anti-scour effect is not up to standard can be analyzed using real-time monitoring data of the deformation of the surrounding rock in the hydraulic slotting area and microseismic monitoring data. After correcting the mechanical parameters of the surrounding rock and the technical parameters of hydraulic slotting, numerical simulations are performed in groups to re-observe the main control parameters affecting the anti-scour effect of multi-level hydraulic slotting weakening under different geological and mining conditions, and apply them to the subsequent mining of the working face. That is, real-time monitoring data continuously verifies and optimizes the numerical simulation model and the simulation model, corrects key parameters, and continues multi-level hydraulic slotting roof weakening operations to continuously maintain the anti-scour capability of the roadway at the end of the working face.

[0090] The method for weakening and preventing rockburst at the end of the working face involves hydraulically cutting the roof in the dangerous area of ​​the roadway at the end of the working face to create pressure relief grooves, blocking grooves, and far-field dynamic load source fracture expansion grooves. This achieves a combined reduction of the energy that can induce rockburst disasters in three aspects: the generation conditions of dynamic loads, the propagation path of dynamic loads, and the static load of the surrounding rock, thereby improving the rockburst resistance of the working face end roadway. The pressure relief grooves created by hydraulic cutting in the roadway roof can release the static load of the roadway surrounding rock, increase the degree of fragmentation of the hard roof, reduce the roof strength and transfer stress to the deeper rock, and at the same time increase the energy dissipation during the propagation of far-field dynamic loads. The blocking grooves created by hydraulic cutting in the path of dynamic loads can reduce the roof strength, transfer stress to the deeper rock, and block the propagation of far-field dynamic loads to the roadway. The far-field dynamic load source fracture expansion grooves created 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 fracture. This method of weakening the roof at the working face end can effectively improve the anti-rockfall capability of the working face end roadway and significantly reduce the possibility of rockburst disasters in the working face end roadway. It is particularly suitable for solving the rockburst problem in the working face end roadway.

Claims

1. A method for weakening and preventing impact on the top plate of a working face, characterized in that, The anti-scouring capability of the roadway at the working face end is improved by hydraulically cutting grooves in the roof near the dynamic load source, the dynamic load propagation path, and the roadway roof. The specific steps include: Step 1: Obtain mine data and 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 technology parameters of the coal seam and surrounding rock of the target fully mechanized longwall face and its goaf. Step 2: Based on the mine data from Step 1, construct a finite element model of the load-deformation structure of the adjacent working face goaf-coal pillar-roadway-coal seam to be mined in this working face, and conduct numerical simulation of the target working face mining process. Utilize data on periodic fracture and collapse of the roof and top coal during mine production, coal seam mining stress data, and roadway surrounding rock deformation and delamination data to verify and optimize the parameters of this load-deformation finite element model, forming a simulation model that conforms to the actual production conditions and monitoring data of the mine site. Based on the simulation model, conduct numerical simulation of the working face mining in the subsequent mining plan, calculate the roadway areas with high static load and large stress concentration coefficient, and the roof fracture areas of adjacent goaf areas, and identify the dangerous roadway areas with the risk of rockburst. Step 3: Based on the coordinates of the dangerous area of ​​the roadway obtained in Step 2 and the coordinates of the far-field dynamic load source obtained from on-site microseismic monitoring, set hydraulic slotting weakening anti-scour conditions that can meet the anti-scour requirements of the roadway. Hydraulic kerving weakens and prevents erosion, and must meet the following conditions: In the formula: Ω represents the rock mass element; W represents the energy of dynamic load propagating to the surrounding rock of the roadway; U t denoted as ψ, representing the strain energy of the rock volume element; t represents time; x represents the volume potential energy of the rock volume element. j For coordinates; σ j Let x be the coordinate j Stress at a location; in, ψ=U s +U b +U cr In the formula: U s Energy consumed during the compaction stage; U b The energy consumed by the development of secondary cracks, the propagation of primary cracks, and the final fracture to form surface energy; U cr This refers to the energy released during the secondary crack fracture friction process; Volumetric strain energy U t The energy W transmitted from the dynamic load to the surrounding rock of the roadway satisfies the following condition: Where: σ s σ is the static load on the surrounding rock of the tunnel; d The dynamic load on the surrounding rock of the tunnel; Initial static load σ of the surrounding rock of the tunnel 0s With initial dynamic load σ 0d The following conditions must be met: In the formula: H n ρ represents the vertical height of coal and rock strata at different burial depths. n denoted as density of different rock layers; g is the acceleration due to 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 Lamé constant of the rock layer; w is the stress wave frequency of the impact load. in, In the formula: c, k, G, and u are the vertical damping coefficient, vertical spring coefficient, shear modulus, and Poisson's ratio of the rock strata, respectively; r is the equivalent radius of the collapsed rock mass; and ρ is the density of the collapsed rock mass. When hydraulic cuts exist in the surrounding rock of the tunnel, the static load σ of the surrounding rock is... s Dynamic load σ of the surrounding rock of the tunnel d All will experience different degrees of decay, and the decayed values ​​will satisfy the following conditions: In the formula: r0 is the radius of the kerf; r x denoted as , where is the distance between the surrounding rock location and the center of the cut; C is the stress wave attenuation coefficient; x0 is the distance between the cut 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 in Step 2 and the hydraulic slotting weakening and anti-scour conditions in Step 3, firstly, simulated drilling is performed in the dangerous area of ​​the roadway in the simulation model, and simulated hydraulic slotting to form pressure relief slots, blocking slots, and far-field dynamic load source fracture expansion slots. Among them, the pressure relief slots are located in the roof of the roadway in the dangerous area, the blocking slots are located on the line connecting the dangerous area of ​​the roadway and the far-field dynamic load source, and the slotting surface of the blocking slots is perpendicular to the direction of dynamic load propagation. The far-field dynamic load source fracture expansion slots are in the same rock layer as the far-field dynamic load source in the vertical direction, and are located near the far-field dynamic load source in the horizontal direction. Then, the dynamic load generated by roof fracture is simulated at the location of the far-field dynamic load source, and the numerical simulation of the working face mining of the subsequent mining plan is carried out to obtain the optimal hydraulic slotting weakening and anti-scour technology parameters that can weaken the dynamic load and reduce the static load. Step 5: Based on the optimized hydraulic slotting weakening and anti-scour technology parameters obtained in Step 4, drilling rigs are deployed in the roadway of the target working face of the mine to drill holes into the roof of the roadway surrounding rock. High-pressure water jets, sprayed from the nozzles at the end of the drill rods and radially along the drill rods, are used to impact and break the roof rock. Through the rotation and retraction of the drill rods, pressure relief slots, blocking slots, and far-field dynamic source fracture expansion slots are formed. This process is repeated to complete the opening of all pressure relief slots, blocking slots, and far-field dynamic source fracture expansion slots. Step 6: Based on the regional and real-time monitoring results of the surrounding rock of the target roadway in the field monitoring area, determine the risk of rockburst in the target roadway before and after hydraulic cutting, evaluate and verify the anti-scour effect of hydraulic cutting, and for areas where the anti-scour effect does not meet the conditions for weakening anti-scour in Step 3, repeat Step 4 and adjust the optimized hydraulic cutting weakening anti-scour technical parameters obtained in Step 4. Based on the adjusted and optimized optimized hydraulic cutting weakening anti-scour technical parameters, perform the hydraulic cutting construction operation in Step 5 again until the anti-scour effect meets the conditions for weakening anti-scour in Step 3.

2. The method for weakening and preventing impact on the top plate of the working face according to claim 1, characterized in that, In Step 2, when performing numerical simulation of the target longwall face mining process, the following steps are taken: First, mine data including the geological conditions and mining conditions of the coal seam in the target longwall face, as well as the basic mechanical parameters, internal porosity data, and fracture distribution data of the coal body and surrounding rock of the target longwall face and its goaf are imported into the load-deformation finite element model. Then, the vertical stress distribution is calculated to obtain the areas with high roof elastic energy density. The mesh is then refined in the areas with high roof elastic energy density. Finally, numerical simulation of the target longwall face mining process is performed to obtain the stress distribution, deformation characteristics, and periodic collapse data of the roof and top coal in the target longwall face and roadway.

3. The method for weakening and preventing impact on the top plate of the working face according to claim 1, characterized in that, In Step 4, when performing simulated hydraulic cutting in the hazardous area of ​​the roadway in the simulation model, the hydraulic cutting depth must meet the following condition: In the formula: p is the jet pressure; R is the nozzle radius; τ 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 on the top plate of the working face according to claim 1, characterized in that, In Step 4, when simulating the formation of pressure relief grooves by hydraulic cutting, simulated drilling is carried out near the junction of the roof coal and the coal pillar in the roadway, and simulated hydraulic cutting is carried out on the side close to the roadway. The downward extension range of the hydraulic cutting does not exceed the junction of 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 on the top plate of the working face according to claim 1, characterized in that, In Step 4, when simulating the formation of a pressure relief groove through hydraulic cutting, after the top plate is weakened, when the direct roof settlement height is ΔH, the roof pressure in the cutting area is... In the formula: E2 is the direct top elastic modulus; L R1 θ1 is the radius of the pressure relief cut; H2 is the height of the immediate roof; q is the pressure of the immediate roof on the coal pillar. The elevation angle θ1 of the pressure relief slotted drill pipe satisfies the following condition: θ1=arctan[(h1-h0) / l0] In the formula: h1 is the roadway 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 on the top plate of the working face according to claim 1, characterized in that, In Step 4, when obtaining the optimal hydraulic slotting weakening and anti-scour technology parameters that can reduce dynamic load and lower static load, group numerical simulations are performed on different parameters of the hydraulic slotting to obtain the dynamic load attenuation rate, roadway surrounding rock stress-strain change characteristics, and coal pillar area stress transfer characteristics under different hydraulic slotting parameter conditions. The main control parameters of hydraulic slotting anti-scour that affect the dynamic load attenuation rate, roadway surrounding rock stress-strain change characteristics, and coal pillar area stress transfer characteristics are determined and obtained. After verification and model optimization based on the mine field observation results, the optimal hydraulic slotting weakening and anti-scour technology parameters are formed.

7. The method for weakening and preventing impact on the top plate of the working face according to claim 1, characterized in that, In Step 6, when evaluating and verifying the anti-scouring effect of hydraulic slotting, microseismic monitoring is conducted in the roadway after slotting. The frequency of mine-induced tremors affecting the roadway before and after slotting is compared, the energy level of mine-induced tremors before and after slotting is compared, the deformation of the surrounding rock in the roadway is measured, and the deformation of the surrounding rock in the slotted area and the unslotted area is compared. The measured data, theoretical model calculations and numerical simulation calculations are compared and verified to evaluate the effect of far-field dilatation slotting in reducing the elastic energy density of the far-field roof and suppressing the occurrence of mine-induced tremors, the effect of stalling slotting in weakening dynamic loads, and the effect of pressure relief slotting in reducing the static load of the roadway surrounding rock.

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