Method for controlling surrounding rock by hydraulic fracturing of deep horizontal hole and shallow vertical hole in advance area of air lane

Abstract

The application discloses a kind of airway advanced area deep level hole-shallow vertical hole hydraulic fracturing roof cutting pressure release surrounding rock control method, it belongs to coal mining field, its technical key points are in at: by the deep-shallow part of airway roof collaborative pressure release, reduce the deformation of surrounding rock in airway advanced area, before mining in working face, drill field is arranged at roadway opening, use kilometer drilling machine in drill field along the trend of roadway, and carry out regional hydraulic fracturing in the key thick hard rock layer in roof deep part, and weaken thick hard rock layer, use drilling machine along the trend of roadway, and set up several vertical presplitting holes to roof in the side of roadway close to coal pillar, before mining in working face, carry out hydraulic fracturing, and weaken thick hard rock layer in the shallow part of roof, after fracturing, in the advanced area of roadway, with mining in working face, use mining water cutting machine to carry out anchor withdrawal operation to anchor cable in roof rock stratum.

Description

Technical Field

[0001] This invention relates to the field of coal mining, and more specifically, to a method for controlling the surrounding rock by hydraulic fracturing and depressurization of the roof in the advanced area of ​​an accessible roadway using deep horizontal holes and shallow vertical holes. Background Technology

[0002] With the increasing depth and intensity of coal mining, the problem of mine pressure control in goaf roadways is becoming increasingly serious. After the initial excavation is completed and the mining phase begins, the goaf roadway is not only affected by the advance support pressure generated by the advancement of the working face, but also by the lateral support pressure from the adjacent goaf area. Under the combined stress, the deformation rate of the surrounding rock in the advance area of ​​the goaf roadway increases, the cumulative deformation increases, affecting the advancement speed of the working face, significantly increasing the amount of roadway repair work and long-term maintenance costs, and seriously restricting the safe and efficient production of the mine. Especially when the immediate roof is a thick and hard rock layer, not only is the roadway deformation severe, easily leading to overhanging roofs at the ends, but it can also cause excessive gas levels in the upper corner, posing a great safety hazard to workers. Therefore, how to reduce the deformation of the surrounding rock in the advance area of ​​the goaf roadway is of great significance.

[0003] Currently, there are three main methods to reduce the deformation of the surrounding rock in the advance zone of tunnels near the exit: (1) Increasing coal pillar width and improving roadway support strength. By increasing the coal pillar width to a sufficient extent, the roadway is positioned in a stress-reducing zone. At the same time, high-strength support is applied to improve the roadway's own bearing capacity, thereby reducing the deformation of the surrounding rock. Although this method has a simple construction process, it wastes a lot of coal resources, and the support cost increases with the mining depth. It also has poor adaptability to strong mining activities at the working face.

[0004] (2) Grouting reinforcement method. By injecting grouting material into the fissures and fractured areas of the surrounding rock, the loose and fractured rock mass is re-cemented into a whole, restoring or even exceeding the original rock strength, thereby resisting the deformation of the surrounding rock caused by mining. This method can enhance the mechanical properties of the surrounding rock, improve its integrity and bearing capacity, and is also adaptable to various roadway cross-sections, with high flexibility. However, the grouting reinforcement method does not change the stress environment of the roadway, and has problems such as long construction period, high cost, and difficulty in precisely controlling the diffusion of grouting material.

[0005] (3) Roof cutting and stress relief method. By actively cutting off the stress transmission path of the roof and changing the overlying rock structure, the roof on the side of the goaf collapses in time, thereby reducing the high stress concentration of the surrounding rock of the roadway. It mainly includes blasting, liquid carbon dioxide fracturing and hydraulic fracturing.

[0006] Blasting involves drilling holes in the roof strata, placing special explosives inside, and using the energy from the explosion to fracture the rock layers, thereby cutting off critical strata and reducing stress concentration. However, this method easily generates large amounts of CO gas during construction, causing CO levels in underground coal mines to exceed limits, threatening personnel safety and seriously polluting the environment, which does not meet the requirements for green mine construction.

[0007] The liquid carbon dioxide fracturing method involves drilling holes at designed locations in the roof and placing carbon dioxide fracturing devices. The instantaneous vaporization of liquid carbon dioxide causes a rapid expansion in volume, generating high pressure that forces the rock mass to produce numerous fractures, thereby reducing stress concentration in the roadway roof. This method is effective when fracturing relatively soft surrounding rocks such as coal seams, but its effectiveness is generally limited when fracturing thick, hard rock strata.

[0008] Hydraulic fracturing involves drilling holes on one side of the tunnel and applying high-pressure water to create fissures in hard rock strata, thereby weakening the rock mass and reducing the supporting pressure on the tunnel roof. However, this method requires drilling through soft rock strata to reach key hard rock strata, limiting the fracturing range and generally resulting in limited effectiveness.

[0009] In summary, increasing the width of the coal pillar and improving the support strength result in significant waste of coal resources. Grouting reinforcement is a passive method of resisting stress transfer to prevent deformation of the surrounding rock in the roadway. In the roof-cutting and pressure-relief method, blasting uses explosive energy to cut through thick, hard rock layers, generating large amounts of CO, leading to CO exceeding limits and posing a high risk. Liquid carbon dioxide fracturing is costly and has poor fracturing effect. Existing hydraulic fracturing methods have significant limitations in their fracturing range, only able to relieve pressure in localized areas, making it difficult to fully release the stress in the surrounding rock of the roadway, resulting in mediocre pressure relief. Based on these problems, this invention proposes a deep horizontal-shallow vertical hole hydraulic fracturing roof-cutting and pressure-relief method for controlling the surrounding rock in the advanced area of ​​an accessible roadway. This method solves the problem of the small fracturing range of traditional hydraulic fracturing, effectively combining regional and precise pressure relief, and has advantages such as a large pressure relief range, good pressure relief effect, and low subsequent support costs. This method can effectively alleviate the problem of large deformation of the surrounding rock in the advanced area of ​​an accessible roadway and conforms to the concept of green and environmentally friendly mine construction. Summary of the Invention

[0010] This invention provides a method for controlling the surrounding rock of a pre-existing tunnel using deep horizontal holes and shallow vertical holes for hydraulic fracturing and depressurization, aiming to solve the problems of existing technologies.

[0011] The specific proposal of this application is as follows: 1. A method for controlling the surrounding rock deformation in the advance zone of an access roadway using hydraulic fracturing with deep horizontal holes and shallow vertical holes, characterized by reducing the deformation of the surrounding rock in the advance zone of the access roadway by synergistically depressurizing the deep and shallow parts of the roof. The method includes the following steps: S1, collect engineering geological data of the air-supported tunnel and working face; S2, determine the location of the deep, thick, hard rock layer in the roof that needs to be fracturing and the hydraulic fracturing parameters, set up a drilling site and install a kilometer drilling rig at the tunnel opening, drill several horizontal boreholes along the tunnel direction toward the target layer, and carry out regional hydraulic fracturing in the boreholes to weaken the deep, thick, hard rock layer. S3. Determine the location of the shallow, thick, hard rock layer in the roof that needs to be fracturing and the hydraulic fracturing parameters. Before the working face is mined, drill vertical pre-fracturing holes into the roof on the side of the roadway near the coal pillar. In the advanced area of ​​the roadway, hydraulic fracturing is carried out in the holes in a step-forward manner to achieve precise pressure relief of the shallow, thick, hard rock layer in the roof. S4. After the fracturing is completed, the working face begins to be mined. During the mining period, the anchor cables in the roof strata are removed using a mine water cutting machine in the advanced area of ​​the access roadway. S5 monitors and analyzes the changes in the surrounding rock approach in the advanced area of ​​the exposed roadway, the overhang area at the working face end, the gas concentration in the corner, and the resistance of the working face end support before and after the roof is depressurized by deep horizontal holes and shallow vertical holes.

[0012] Furthermore, the deep rock strata of the roof refer to the rock strata within a range of 30 to 60 meters from the roof; the shallow rock strata of the roof refer to the rock strata within a range of 10 to 30 meters from the roof.

[0013] Furthermore, the engineering geological data in step S1 includes: tunnel excavation operation procedures and support cross-section diagrams, working face mining operation procedures, mining engineering plan, surrounding borehole columnar section diagrams, geological report, and mine preliminary design specification.

[0014] Furthermore, step S2 includes the following sub-steps: S2.1 Based on the data collected in step S1 and combined with the actual situation on site, determine the key rock layers and their thicknesses that require hydraulic fracturing for the kilometer-long horizontal borehole. S2.2, by taking core samples in the field and conducting laboratory mechanical tests, the compressive strength and tensile strength of the target fracturing rock layer are obtained, and the injection pressure required for hydraulic fracturing is obtained accordingly; S2.3, Calculate the effective fracture propagation radius R of the borehole based on the hydraulic fracturing parameters; R = 0.8546 × [π] 0.5 ·E·Q·t / (K I ·(1-v 2 )·4×2 0.5 )] 0.4 In the formula: E—Rock elastic modulus, in MPa; Q—hydraulic fracturing fluid injection rate, in L / min; t—fracturing time, in minutes; K I—Rock fracture propagation toughness; v—Poisson's ratio; S2.4. Based on the effective fracture propagation radius of the borehole, ensure that the fracturing range of the horizontal borehole can cover the target thick and hard rock layer, thereby determining the number of horizontal boreholes to be drilled per kilometer and the drilling location. S2.5, set up a drilling site near the tunnel opening, install a kilometer drilling rig and a high-pressure water injection pump in the drilling site, drill horizontal long holes into the target thick and hard rock layer according to the design parameters, and then start hydraulic fracturing. Each hole adopts the retreating multiple fracturing method, and fracturing is carried out every 10~30m, with each fracturing time not less than 25min. S2.6 During fracturing, as the fracturing time increases, the water pressure gradually rises to its maximum value. Due to the high pressure, new cracks appear in the fracturing rock layer and begin to expand. As the water pressure decreases slightly, the cracks continue to spread outwards. When the water pressure decreases significantly, the fracturing is complete.

[0015] Furthermore, step S3 includes the following sub-steps: S3.1, Determine the cutting height of the vertical pre-splitting hole; Based on the relevant hydraulic fracturing parameters of the horizontal borehole, the cut-off height of the vertical pre-fracturing borehole is calculated according to the following formula: H v =H h -W f / 2-h s ; In the formula: H v —Vertical hole hydraulic fracturing cut-off height (m); H h —Horizontal hole hydraulic fracturing cut-off height (m); W f — Horizontal hole fracturing fracture propagation length (m); h s —Thickness of the intermediate soft rock layer between the horizontal hole decompression zone and the vertical hole decompression zone (m); This formula is based on the spatial connection between the horizontal and vertical borehole fracturing regions. Both horizontal and vertical borehole fracturing fractures extend outwards from the borehole center, but their primary directions of extension differ. Horizontal borehole fractures primarily extend vertically; therefore, the decompression range of a horizontal borehole is between H... h -W f / 2 and H h +W f Between / 2, and the fractures in the vertical holes mainly extend horizontally. To ensure effective connection between the shallow and deep fracturing zones, the theoretical cut-off height of the vertical holes should be H. v =H h -W f / 2, further considering the presence of soft rock between the deep and shallow fracturing zones, which has poor integrity and can naturally collapse under mining pressure and its own weight, the thickness of the soft rock should be subtracted to avoid over-construction. Therefore, the final formula is H. v =H h -W f / 2-h s ; S3.2, Determine the hydraulic fracturing parameters. Based on the cut-off height of the vertical pre-fracturing hole, the compressive and tensile strengths of the rock layer to be fractured are obtained through on-site core sampling and laboratory tests, thereby determining the required water injection pressure; S3.3, Calculate the effective crack propagation radius R of the borehole based on the obtained parameters, and determine the borehole spacing D: D=2R; R = 0.8546 × [π] 0.5 ·E·Q·t / (K I ·(1-v 2 )·4×2 0.5 )] 0.4 In the formula: E—Rock elastic modulus, in MPa; Q—hydraulic fracturing fluid injection rate, in L / min; t—fracturing time, in minutes; K I —Rock fracture propagation toughness; v—Poisson's ratio; S3.4 Using a drilling rig, according to the determined parameters of the vertical pre-splitting holes, drill holes vertically towards the roof of the roadway along the roadway direction, close to the coal pillar side. S3.5 After drilling is completed, install a high-pressure water pump and perform vertical hole hydraulic fracturing in a step-forward manner within 20m ahead of the access roadway. The fracturing time shall not be less than 15 minutes. When the water pressure drops significantly or a large amount of water flows out of the adjacent hole, the fracturing is completed.

[0016] Furthermore, step S4 includes: after fracturing is completed, a water-jet cutting machine is installed. As the working face begins to be mined, the anchor cables of the roof within 1m ahead of the roadway are water-jet cut. The cutting head is precisely positioned at the predetermined cutting point of the anchor cable to be cut, ensuring that the cutting head is perpendicular to the anchor cable. The cutting sequence is carried out row by row along the roadway direction.

[0017] Furthermore, step S5 includes: through on-site monitoring, comparing the amount of surrounding rock approaching the advance area of ​​the roadway before and after the coordinated pressure relief control of deep horizontal holes and shallow vertical holes on the roof, the overhang area at the working face end, the gas concentration in the upper corner, and the resistance of the support at the working face end.

[0018] The beneficial effects of this application are as follows: First, this invention proposes a method for controlling the pressure relief of surrounding rock in the advanced area of ​​an access roadway using hydraulic fracturing with deep horizontal holes and shallow vertical holes, spatially constructing a synergistic pressure relief system for the deep and shallow sections. Specifically, a kilometer-long drilling rig is used to drill horizontal holes in deep, thick, hard rock strata for hydraulic fracturing, achieving regional pressure relief of the deep surrounding rock in the access roadway. Vertical pre-fracturing holes are drilled in the advanced area of ​​the roadway for hydraulic fracturing, weakening the shallow, thick, hard rock strata and achieving precise directional pressure relief of the shallow surrounding rock. Then, water jet anchor removal is used to weaken the anchoring force of the roof anchor cables, making the roof more prone to collapse after the working face is pushed over. These three methods act spatially on both the deep and shallow parts of the surrounding rock, combining regional and precise pressure relief to form a comprehensive three-dimensional pressure relief system from the far field to the near field, effectively solving the problems of small pressure relief range and poor effect of traditional hydraulic fracturing.

[0019] Second, this invention determines a cutting height H for a vertical pre-splitting hole. v Calculation method: H v =H h -W f / 2-h s ; A method was established to determine the relationship between the top cutting height of a kilometer-long horizontal borehole and the top cutting height of a vertical pre-splitting borehole. Based on the actual fracturing effect of the horizontal borehole, the top cutting height of the vertical pre-splitting borehole was rationally determined to ensure that the deep fracturing zone formed by the horizontal borehole and the shallow fracturing zone formed by the vertical borehole are effectively connected in space, significantly improving the surrounding rock pressure relief control effect in the advanced area of ​​the access roadway.

[0020] Third, this invention establishes a temporal connection between three elements: deep horizontal boreholes for regional stress relief, shallow vertical boreholes for precise local stress relief, and surface refinement and anchor removal. Specifically, during tunnel excavation, a drilling site can be pre-drilled, and horizontal boreholes can be constructed and hydraulically fracturing performed using a kilometer-long drilling rig. This weakens the deep, thick, and hard rock layers in advance, altering the stress environment of the surrounding rock. Before the working face begins mining, vertical pre-fracturing boreholes are drilled in the advanced area of ​​the tunnel and hydraulically fracturing is performed. Based on the weakening effect achieved through the horizontal boreholes, the stress transmission path of the shallow rock layers is further precisely cut off. When mining begins, as the working face advances, the anchor cable constraint on the roof is released through water jet anchor removal operations, preventing the formation of a suspended roof at the working face end. These three elements form a mutually progressive and coordinated relationship in time, significantly reducing stress concentration in the surrounding rock of the tunnel and improving construction efficiency as well as the stability and safety of the surrounding rock in tunnels near the exit. Attached Figure Description

[0021] Figure 1 This is a design schematic diagram of Example 1 (without hydraulic fracturing).

[0022] Figure 2 This is a design schematic diagram of Example 1 (conducting hydraulic fracturing).

[0023] Figure 3 yes Figure 2 AA section view.

[0024] Figure 4 yes Figure 2 BB cross-section diagram.

[0025] Figure 5 This is a design schematic diagram of Example 2 (without hydraulic fracturing).

[0026] Figure 6 This is a design schematic diagram of Example 2 (conducting hydraulic fracturing).

[0027] Figure 7 yes Figure 6 AA section view.

[0028] Figure 8 yes Figure 6 BB cross-section diagram. Detailed Implementation

[0029] To better understand the above-mentioned objectives, features, and advantages of the present invention, the present invention will be further described in detail below with reference to the accompanying drawings and specific embodiments. It should be noted that, unless otherwise specified, the embodiments and features described in these embodiments can be combined with each other.

[0030] Many specific details are set forth in the following description in order to provide a full understanding of the invention. However, the invention may also be practiced in other ways different from those described herein, and therefore the scope of protection of the invention is not limited to the specific embodiments disclosed below.

[0031] Example 1: The 10105 longwall mining face in a certain mine As shown in the figure, the present invention provides a method for controlling the surrounding rock pressure by hydraulic fracturing and roof-cutting in the advanced area of ​​an open roadway. The specific method is carried out according to the following steps: Step I: Collect engineering geological data for the open roadways and working faces on-site, such as roadway excavation operation procedures, support cross-section diagrams, working face mining operation procedures, mining engineering plan, surrounding borehole columnar sections, geological reports, and preliminary mine design specifications. Use a sighting instrument to inspect the roof and record the roof lithology at different locations. Based on the collected data, it is known that: the 10105 working face mines the No. 10 coal seam, with an average thickness of 5.7m, using fully mechanized caving mining, a machine mining height of 3m, a top coal caving height of 2.7m, and a mining-to-caving ratio of 1:0.9. The 10105 transport roadway is 1500m long, with a tunneling width of 5.5m and a tunneling height of 3.5m. The 10105 cut-off section is designed to be 200m long, with a cross-sectional width of 7.7m and a height of 3m. The coal mine is a high-gas mine. The roof support uses Φ20×2400mm threaded steel anchor bolts, and the top anchor cables use Φ21.6mm×6500mm low-relaxation steel strands. The roof consists of, in sequence, No. 10 coal seam (2.7m thick), fine-grained sandstone (7.2m thick), medium-grained sandstone (4m thick), mudstone (2m thick), sandy mudstone (4m thick), limestone (5.5m thick), sandy mudstone (2.5m thick), mudstone (3.6m thick), No. 7 coal seam (0.4m thick), limestone (7m thick), siltstone (3m thick), sandy mudstone (2.9m thick), mudstone (2.6m thick), limestone (3.5m thick), sandy mudstone (5m thick), mudstone (4m thick), medium-grained sandstone (4.5m thick), and sandy mudstone (5.5m thick), etc. See details below. Figure 3 .

[0032] Step II: Based on the exploration and observation results, the key thick and hard rock layers in the deep area were identified as a 7m thick limestone layer (31.9m from the top plate), a 3.5m thick limestone layer (47.4m from the top plate), and a 4.5m thick medium-grained sandstone layer (59.9m from the top plate). The elastic moduli of the 7m and 3.5m thick limestone layers are 11.7GPa and 12.3GPa, respectively, and their Poisson's ratios are 0.21 and 0.24, respectively. The calculated hydraulic fracturing fracture propagation radii are 3.5m and 3.7m, respectively. The elastic modulus of the medium-grained sandstone is 6.6GPa, and its Poisson's ratio is 0.22, resulting in a hydraulic fracturing fracture propagation radius of 3.1m. Therefore, a total of four horizontal boreholes are required. Two horizontal boreholes are needed for the 7m thick limestone layer, and one horizontal borehole is needed for each of the 3.5m thick limestone layer and the 4.5m thick medium-grained sandstone layer. According to laboratory mechanical tests, the fracturing time for horizontal hole hydraulic fracturing is 25 minutes.

[0033] Step III: Before the working face is mined, a drilling site is set up at the roadway opening. A kilometer-long drilling rig is installed in the drilling site, followed by a high-pressure water injection pump. Hydraulic fracturing then begins. The initial part of the borehole trajectory is a curve, while the main part is an approximately horizontal straight line.

[0034] Step IV: Based on the inspection results, the soft rock thickness below the horizontally fractured rock layer is 6.5m. Therefore, the calculated cut-off height of the vertical pre-splitting hole is 25.4m, and 25m is adopted for construction. Based on the drilling depth and the mechanical properties of the rock, the hydraulic fracturing time for the vertical pre-splitting hole is determined to be 15 minutes. The thick, hard rock layer to be fracturing in the vertical hole is limestone, with an elastic modulus of 12.1 GPa and a Poisson's ratio of 0.22. The calculated vertical hole spacing is 8.3m. To ensure fracturing effect and for convenience on site, the vertical pre-splitting hole spacing is adopted to be 8m for construction.

[0035] Step V: Drill vertical pre-splitting holes on the side close to the coal pillar according to the determined parameters. Before the working face is mined, hydraulic fracturing is carried out within a range of 20m ahead. After the fracturing is completed, the roof anchor cables are cut and unanchored within a range of 1m ahead of the working face using a mining water cutting machine. Then mining begins.

[0036] Step VI: As the working face advances, repeat Step V until the working face mining is completed. During mining, monitor the resistance of the hydraulic support at the working face end. Compare and analyze the approach of the surrounding rock in the advanced area of ​​the roadway before and after fracturing, the overhanging roof area at the working face end, and the gas concentration in the corner. Through monitoring and calculation, it is found that after combined hydraulic fracturing, the maximum approach of the roof and floor decreased from 953mm to 357mm, the maximum approach of the two sides decreased from 893mm to 303mm, and the overhanging roof area at the end decreased from 18m². 2 Reduced to 9.8m 2 The gas concentration in the upper corner of the working face decreased from 0.73% to about 0.47%, and the working resistance of the end support decreased by a maximum of 21.7%.

[0037] Example 2: 8104 longwall face of a certain No. 2 mine As shown in the figure, the present invention provides a method for controlling the surrounding rock pressure by hydraulic fracturing and roof-cutting in the advanced area of ​​an open roadway. The specific method is carried out according to the following steps: Step 1: Collect engineering geological data for the roadways and working faces on-site, such as roadway excavation operation procedures, support cross-section diagrams, working face mining operation procedures, mining engineering plan, surrounding borehole columnar sections, geological reports, and preliminary mine design specifications. Use a sighting instrument to inspect the roof and record the roof lithology at different locations. Based on the collected data, the average thickness of coal seam No. 8 is 2.5m. A longwall retreat mining method with full-height mechanized mining is adopted, and the roof management uses the full caving method. The 8104 return airway is 1400m long, 5m wide, and 2.8m high. The designed length of the 8104 cut-off is 155m, with a cross-section of 7.2m wide and 2.8m high. The mine is a low-gas mine. The roof support uses high-strength threaded steel anchor bolts with a specification of Φ18×2000mm, and the anchor cables are 7-strand steel strands with a specification of Φ17.8×6200mm. The roof consists of limestone (3.2m thick), siltstone (1.8m thick), limestone (6.6m thick), siltstone (6.9m thick), limestone (10m thick), sandy mudstone (4.7m thick), medium sandstone (7.7m thick), sandy mudstone (4.2m thick), siltstone (8.3m thick), and sandy mudstone (7m thick), etc., as detailed in the attached document. Figure 7 .

[0038] Step II: Based on the exploration and observation results, the key hard rock layers in the deep area were identified as a 7.7m thick medium-grained sandstone (33.2m from the top plate) and an 8.3m thick siltstone (45.1m from the top plate). The elastic moduli of the medium-grained sandstone and siltstone are 8.3GPa and 11.1GPa, respectively, and their Poisson's ratios are 0.23 and 0.26, respectively. The calculated fracture propagation radii for hydraulic fracturing are 3.2m and 3.5m, respectively, therefore a total of 4 horizontal boreholes are required. Two horizontal boreholes are needed for the 7.7m thick medium-grained sandstone, and two boreholes are needed for the 8.3m thick siltstone. According to laboratory mechanical tests, the fracturing time for horizontal borehole hydraulic fracturing is 30 minutes.

[0039] Step III: Before the working face is mined, a drilling site is set up at the roadway opening. A kilometer-long drilling rig is installed in the drilling site, followed by a high-pressure water injection pump. Hydraulic fracturing then begins. The initial part of the borehole trajectory is a curve, while the main part is an approximately horizontal straight line.

[0040] Step IV: Based on the inspection results, the soft rock thickness below the horizontally fractured rock layer is 4.7m. Therefore, the calculated cut-off height of the vertical pre-splitting hole is 28.1m, and 28m is adopted for construction. Based on the drilling depth and the mechanical properties of the rock, the hydraulic fracturing time for the vertical pre-splitting hole is determined to be 15 minutes. The thick, hard rock layer to be fractured by the vertical hole is limestone, with an elastic modulus of 7.1 GPa and a Poisson's ratio of 0.2. The calculated vertical hole spacing is 8.7m. To ensure fracturing effect and for convenience on site, the vertical pre-splitting hole spacing is adopted as 8.5m for construction.

[0041] Step V: Drill vertical pre-splitting holes on the side close to the coal pillar according to the determined parameters. Before the working face is mined, hydraulic fracturing is carried out within a range of 20m ahead. After the fracturing is completed, the roof anchor cables are cut and unanchored within a range of 1m ahead of the working face using a mining water cutting machine. Then mining begins.

[0042] Step VI: As the working face advances, repeat Step V until the working face mining is completed. During mining, monitor the resistance of the hydraulic support at the working face end. Compare and analyze the advancing rate and approach amount of the surrounding rock in the advance zone of the roadway before and after fracturing, as well as the overhanging roof area at the working face end. Through monitoring and calculation, it is found that after combined hydraulic fracturing, the maximum approach amount of the roof and floor decreased from 871mm to 339mm, the maximum approach amount of the two sidewalls decreased from 758mm to 294mm, and the overhanging roof area at the end decreased from 41.0m². 2 Reduced to 15.3m 2 The working resistance of the end support was reduced by up to 22.3%.

[0043] The above-described embodiments are preferred embodiments of the present invention and are only used to facilitate the illustration of the present invention. They are not intended to limit the present invention in any way. Any person skilled in the art who makes local modifications or alterations to the technical content disclosed in the present invention without departing from the scope of the technical features of the present invention shall still fall within the scope of the technical features of the present invention.

Claims

1. A method for controlling surrounding rock pressure during hydraulic fracturing with deep horizontal holes and shallow vertical holes in the advance zone of an open roadway, characterized in that... By synergistically relieving pressure on the deep and shallow sections of the roof of the access roadway, the deformation of the surrounding rock in the advance zone of the access roadway is reduced, including the following steps: S1, collect engineering geological data of the air-supported tunnel and working face; S2, determine the location of the deep, thick, hard rock layer in the roof that needs to be fracturing and the hydraulic fracturing parameters, set up a drilling site and install a kilometer drilling rig at the tunnel opening, drill several horizontal boreholes along the tunnel direction toward the target layer, and carry out regional hydraulic fracturing in the boreholes to weaken the deep, thick, hard rock layer. S3. Determine the location of the shallow, thick, hard rock layer in the roof that needs to be fracturing and the hydraulic fracturing parameters. Before the working face is mined, drill vertical pre-fracturing holes into the roof on the side of the roadway near the coal pillar. In the advanced area of ​​the roadway, hydraulic fracturing is carried out in the holes in a step-forward manner to achieve precise pressure relief of the shallow, thick, hard rock layer in the roof. S4. After the fracturing is completed, the working face begins to be mined. During the mining period, the anchor cables in the roof strata are removed using a mine water cutting machine in the advanced area of ​​the access roadway. S5 monitors and analyzes the changes in the surrounding rock approach in the advanced area of ​​the exposed roadway, the overhang area at the working face end, the gas concentration in the corner, and the resistance of the working face end support before and after the roof is depressurized by deep horizontal holes and shallow vertical holes.

2. The method for controlling surrounding rock pressure by hydraulic fracturing with deep horizontal holes and shallow vertical holes in the advanced area of ​​an open roadway according to claim 1, characterized in that, The deep rock strata of the roof refer to the rock strata within a range of 30 to 60 meters from the roof; the shallow rock strata of the roof refer to the rock strata within a range of 10 to 30 meters from the roof.

3. The method for controlling surrounding rock pressure by hydraulic fracturing with deep horizontal holes and shallow vertical holes in the advanced area of ​​an open roadway according to claim 1, characterized in that, The engineering geological data for step S1 includes: tunnel excavation operation procedures and support cross-section diagrams, working face mining operation procedures, mining engineering plan, surrounding borehole columnar section diagrams, geological report, and mine preliminary design specification.

4. The method for controlling surrounding rock pressure by hydraulic fracturing with deep horizontal holes and shallow vertical holes in the advanced area of ​​an open roadway according to claim 1, characterized in that, Step S2 includes the following sub-steps: S2.1 Based on the data collected in step S1 and combined with the actual situation on site, determine the key rock layers and their thicknesses that require hydraulic fracturing for the kilometer-long horizontal borehole. S2.2, by taking core samples in the field and conducting laboratory mechanical tests, the compressive strength and tensile strength of the target fracturing rock layer are obtained, and the injection pressure required for hydraulic fracturing is obtained accordingly; S2.3, Calculate the effective fracture propagation radius R of the borehole based on the hydraulic fracturing parameters; R=0.8546×[π 0.5 ·E·Q·t / (K I ·(1-v 2 )·4×2 0.5 )] 0.4 In the formula: E—Rock elastic modulus, in MPa; Q—hydraulic fracturing fluid injection rate, in L / min; t—fracturing time, in minutes; K I —Rock fracture propagation toughness; v—Poisson's ratio; S2.

4. Based on the effective fracture propagation radius of the borehole, ensure that the fracturing range of the horizontal borehole can cover the target thick and hard rock layer, thereby determining the number of horizontal boreholes to be drilled per kilometer and the drilling location. S2.5, set up a drilling site near the tunnel opening, install a kilometer drilling rig and a high-pressure water injection pump in the drilling site, drill horizontal long holes into the target thick and hard rock layer according to the design parameters, and then start hydraulic fracturing. Each hole adopts the retreating multiple fracturing method, and fracturing is carried out every 10~30m, with each fracturing time not less than 25min. S2.6 During fracturing, as the fracturing time increases, the water pressure gradually rises to its maximum value. Due to the high pressure, new cracks appear in the fracturing rock layer and begin to expand. As the water pressure decreases slightly, the cracks continue to spread outwards. When the water pressure decreases significantly, the fracturing is complete.

5. The method for controlling surrounding rock pressure by hydraulic fracturing with deep horizontal holes and shallow vertical holes in the advanced area of ​​an open roadway according to claim 1, characterized in that, Step S3 includes the following sub-steps: S3.1, Determine the cutting height of the vertical pre-splitting hole; Based on the relevant hydraulic fracturing parameters of the horizontal borehole, the cut-off height of the vertical pre-fracturing borehole is calculated according to the following formula: H v =H h -W f / 2-h s ; In the formula: H v —Vertical hole hydraulic fracturing cut-off height (m); H h —Horizontal hole hydraulic fracturing cut-off height (m); W f — Horizontal hole fracturing fracture propagation length (m); h s —Thickness of the intermediate soft rock layer between the horizontal hole decompression zone and the vertical hole decompression zone (m); S3.2, Determine the hydraulic fracturing parameters. Based on the cut-off height of the vertical pre-fracturing hole, the compressive and tensile strengths of the rock layer to be fractured are obtained through on-site core sampling and laboratory tests, thereby determining the required water injection pressure; S3.3, Calculate the effective crack propagation radius R of the borehole based on the obtained parameters, and determine the borehole spacing D: D=2R; R=0.8546×[π 0.5 ·E·Q·t / (K I ·(1-v 2 )·4×2 0.5 )] 0.4 In the formula: E—Rock elastic modulus, in MPa; Q—hydraulic fracturing fluid injection rate, in L / min; t—fracturing time, in minutes; K I —Rock fracture propagation toughness; v—Poisson's ratio; S3.4 Using a drilling rig, according to the determined parameters of the vertical pre-splitting holes, drill holes vertically towards the roof of the roadway along the roadway direction, close to the coal pillar side. S3.5 After drilling is completed, install a high-pressure water pump and perform vertical hole hydraulic fracturing in a step-forward manner within 20m ahead of the access roadway. The fracturing time shall not be less than 15 minutes. When the water pressure drops significantly or a large amount of water flows out of the adjacent hole, the fracturing is completed.

6. The method for controlling surrounding rock pressure by hydraulic fracturing with deep horizontal holes and shallow vertical holes in the advanced area of ​​an open roadway according to claim 1, characterized in that, Step S4 includes: After fracturing is completed, a water-jet cutting machine is installed. As the working face begins to be mined, the anchor cables of the roof within 1m ahead of the roadway are water-jet cut. The cutting head is precisely positioned at the predetermined cutting point of the anchor cable to be cut, ensuring that the cutting head is perpendicular to the anchor cable. The cutting sequence is carried out row by row along the roadway direction.

7. The method for controlling surrounding rock pressure by hydraulic fracturing with deep horizontal holes and shallow vertical holes in the advanced area of ​​an open roadway according to claim 1, characterized in that, Step S5 includes: through on-site monitoring, comparing the amount of surrounding rock approaching the advance area of ​​the roadway before and after the coordinated pressure relief control of deep horizontal holes and shallow vertical holes on the roof, the overhang area at the working face end, the gas concentration in the upper corner, and the resistance of the support at the working face end.

Images