Method for treating end suspension top through cooperation of frosted water jet grooving and hydraulic fracturing layering top cutting

Abstract

The invention belongs to the technical field of coal mining, and provides a method for treating an end suspended roof through cooperation of frosted water jet grooving and hydraulic fracturing layering roof cutting. A drilling machine is used for drilling holes in a hard roof of a target rock stratum on the two sides of a roadway, namely the coal pillar side and the stoping side at the same time, and annular cutting grooves parallel to a roof layer are cut in different layers of the target rock stratum through the drilled holes within the advance range of a working face by means of an abrasive water jet type coal seam slotting pressure relief device. And hydrofracture is carried out at the position of the cutting groove, and the thick hard top plate is divided into multiple layers of thin hard top plates. After the working face is pushed, the top plate is easier to collapse, the end suspended roof area of the working face is reduced, the end support working resistance is reduced, corner gas overrun is avoided, and the roadway stability is improved. In the construction process, safety is high, the working environment is good, relevant parameters can be optimized and adjusted according to peeping results, and the fracturing effect is good. According to the method, the fracturing effect can be effectively improved, and the method conforms to the green and environment-friendly mine construction concept.

Description

Technical Field

[0001] The present invention belongs to the technical field of coal mining, and particularly relates to a method for treating the end suspended roof by combining abrasive water jet grooving with hydraulic fracturing for layered roof cutting. Background Technique

[0002] After the open-off cut in the coal mine underground is excavated and enters the coal mining stage, when the working face is mined, the roof of the roadway is a hard and thick rock stratum that is not easy to collapse, forming a large area of suspended roof in the goaf. Especially, the sudden collapse of the large area of suspended roof in the end triangular area will squeeze out the gas in the goaf, resulting in over-limit gas in the corner, posing a great safety hazard to the staff. Therefore, shortening the suspended roof at the end of the working face is of great significance.

[0003] At present, there are mainly the following three methods for shortening the length of the suspended roof at the end of the working face: (1) Blasting method. By drilling holes in the key stratum of the roadway roof and setting blasting agents in the holes, the explosion energy is used to cause the rock stratum to rupture, so as to reduce the end suspended roof. In this method, a large amount of gas is easily accumulated in the end suspended roof area, and safety accidents are more likely to occur during the blasting roof caving process. Moreover, it is easily affected by safety control, and the supply of blasting materials is often interrupted, unable to ensure the orderly progress of normal engineering construction.

[0004] (2) Hydraulic fracturing method. By drilling vertical holes on one side of the roadway and applying high-pressure water to cause cracks in the hard rock stratum, the roof pressure is reduced, and the area of the end suspended roof is reduced. The effect of this method is generally average.

[0005] (3) Liquid carbon dioxide pre-fracturing method. This method is safe, environmentally friendly, does not pollute the surrounding environment, has a small seismic velocity, and the transportation, storage and use of liquid carbon dioxide do not require approval. However, the effect of generating cracks by this method is generally average, and the cost is high. Summary of the Invention

[0006] In order to solve the above problems, the present invention provides a method for treating the end suspended roof by combining abrasive water jet grooving with hydraulic fracturing for layered roof cutting.

[0007] The present invention is realized by the following technical solutions: A method for treating the end suspended roof by combining abrasive water jet grooving with hydraulic fracturing for layered roof cutting. Using a drill rig to drill holes in the hard roof of the target stratum simultaneously on both sides of the roadway, namely the coal pillar side and the mining side. In the advance range of the working face, an abrasive water jet coal seam cutting and pressure relief device is used to cut circumferential grooves parallel to the roof plane at different layers of the target stratum through the holes, and hydraulic fracturing is carried out at the grooving position to divide the thick and hard roof into multiple thin and hard roofs.

[0008] Further, it includes the following steps: S1, collecting the engineering geological data of the roadway and the working face; S2. Determine the drilling angle, drilling spacing, circumferential grooving, and water pressure fracturing parameters; S3. Drill holes along the advancing direction of the working face towards the roof according to the design parameters until the target horizon is reached, obtaining multiple parallel hydraulic pressure relief holes; S4. In the advanced area of the working face, use an abrasive water jet coal seam slotting pressure relief device to conduct circumferential grooving in the drill holes and carry out hydraulic fracturing at the grooving position; S5. After the water pressure fracturing is completed, use a peephole to analyze the fracturing effect. Then, the working face starts to be mined, and specifically analyze the suspended roof area at the end of the working face, the gas concentration at the corner, and the stress on the end support before and after roof cutting.

[0009] Furthermore, in step S1, the geological data includes: the roadway driving operation regulations and support section drawings, the mining and excavation engineering plan, the surrounding drill hole columnar diagrams, the geological report, and the preliminary design specifications of the mine.

[0010] In step S2, through theoretical calculation, on-site peeping, and laboratory tests, determine the drilling angle, drilling spacing, target horizon thickness, and the required circumferential grooving position; specifically including the following steps: S2.1 Determine the drilling spacing: The drilling spacing is 2 times the crack propagation radius of the hydraulic fracturing, specifically: D = 2 R ; where: R — Crack propagation radius, m; D — Drilling spacing, m; The crack propagation radius of the hydraulic fracturing is based on the following formula: ; where: Q — Injection rate of the hydraulic fracturing fluid, L / min; E — Elastic modulus of the rock, MPa; σ y — Rock fracture strength, MPa; K I — Rock crack propagation toughness; S2.2. Determine the drilling angle: Before the working face is mined, according to the designed hole spacing, combined with the occurrence of the suspended roof at the end during the mining process of the working face under similar engineering geological conditions, use a drill rig to drill holes vertically to the roof on the side close to the coal pillar, and drill the holes to the midpoint position of the side triangle parallel to the starting cut eye on the side close to the mining side. Specifically, there are the following formulas: ; ; where: α — Drilling azimuth angle, °; β — Drilling inclination angle, °; D — Drilling spacing, m; y — Length of the side triangle parallel to the starting cut eye, m; h — Vertical height from the fracturing horizon to the roof (m); Calculated according to the above formula: When the fracturing interval is located in conglomerate or limestone with uniaxial compressive strength greater than 60 MPa, the crack propagation toughness is taken as 1.2; the borehole azimuth α: ; the borehole dip angle β : ; When the fracturing interval is located in fine-grained sandstone or medium-grained sandstone with uniaxial compressive strength ranging from 50 to 60 MPa, the crack propagation toughness is taken as 1.0; the borehole azimuth α: ; the borehole dip angle β: ; When the fracturing interval is located in fine-grained sandstone or medium-grained sandstone with uniaxial compressive strength ranging from 40 to 50 MPa, the crack propagation toughness is taken as 0.8; the borehole azimuth α: ; the borehole dip angle β: ; S2.3. Determine the number and position of the holes according to the thickness of the hard rock layer at the target interval: When the thickness of the hard rock layer is less than 5 m, prefabricate 1 circumferential groove in the borehole and the groove is set at the 1 / 2 position of the rock layer; when the thickness of the cut hard rock layer is 5 - 10 m, prefabricate 2 circumferential grooves in the borehole and the grooves are set at the 1 / 3 position of the rock layer; when the thickness of the cut hard rock layer is 10 - 15 m, prefabricate 3 circumferential grooves in the borehole and each groove is successively set at the positions of 1 / 4, 1 / 2, and 3 / 4 from the boundary of the rock layer; S2.4. Obtain the required abrasive water jet flow rate, abrasive water jet pressure, abrasive concentration, and water injection pressure by testing the compressive and tensile strengths of the required fractured thick rock layer in the laboratory.

[0011] Furthermore, step S3 is to use a drilling rig to drill holes along the advancing direction of the working face towards the roof on both sides of the roadway, namely the coal pillar side and the mining side, according to the borehole design parameters in S2 until reaching the target interval.

[0012] Furthermore, step S4 includes the following steps: S4.1. Withdraw the drilling rig, insert the abrasive water jet coal seam slotting and pressure relief device into the drilled hole, rotate the drill pipe at a constant speed, and conduct circumferential grooving according to the design parameters in S2; S4.2. Start hydraulic fracturing using the high-pressure water pump of the abrasive water jet coal seam slotting and pressure relief device. Each fracturing time is ≥ 30 minutes, and the water pressure drops significantly, at which point the fracturing is completed. Due to the high pressure, new cracks appear in the fractured rock layer and begin to expand. The water pressure drops slightly, and the cracks spread in all directions. Then the water pressure drops significantly, and the fracturing is completed.

[0013] In step S5: After the hydraulic fracturing is completed, the borehole is peeped again using a peephole to analyze the hydraulic fracturing effect; after the hydraulic fracturing is completed, the working face starts to be mined, and the cantilever roof area, corner gas concentration, and support stress conditions at the working face end are compared and analyzed before and after roof cutting.

[0014] The present invention proposes to combine abrasive water jet circumferential slotting with hydraulic fracturing for layered roof cutting. Using a drilling rig, boreholes are drilled into the hard roof of the target rock layer on both sides of the roadway, i.e., the coal pillar side and the mining side, simultaneously. Within the advanced range of the working face, an abrasive water jet coal seam cutting pressure relief device is used to cut circumferential slots parallel to the roof plane at different layers of the hard roof of the target rock layer through the boreholes and perform hydraulic fracturing at the slotting positions, thereby dividing the thick and hard roof into multiple thin and hard roofs. After the working face has passed, the roof is more likely to collapse, reducing the cantilever roof area at the working face end, lowering the working resistance of the end supports, avoiding corner gas overrun, and improving the roadway stability. The abrasive water jet roof layered cutting technology is adopted, which has high safety during the construction process, a good working environment, can optimize and adjust relevant parameters according to peeping results, and has a good fracturing effect. This method can effectively improve the fracturing effect and conforms to the concept of green and environmental-friendly mine construction.

[0015] The beneficial effects of this application are as follows: First, the present invention uses a drilling rig to drill boreholes into the hard rock layer of the target layer on both sides of the roadway, i.e., the coal pillar side and the mining side, simultaneously. Within the advanced range of the working face, an abrasive water jet coal seam cutting pressure relief device is used to cut circumferential slots parallel to the roof plane at different layers of the hard rock layer through the boreholes and perform hydraulic fracturing at the slotting positions, thereby dividing the thick and hard roof into multiple thin and hard roofs. After the working face has passed, the roof is more likely to collapse, reducing the cantilever roof area at the working face end, avoiding corner gas overrun, and improving the roadway stability.

[0016] Second, the present invention effectively reduces the cantilever roof length and improves the stability of the extraction roadway by drilling a borehole perpendicular to the roof near the coal pillar side and drilling a borehole to the midpoint of the side parallel to the starting cut line of the side triangle near the mining side. Since the cantilever roof is similar to a triangular plate structure and has high stability itself and is not easily damaged as a whole, after hydraulic fracturing at the midpoint of the side line, cracks expand from the midpoint of the side line to both sides, cutting off the side line parallel to the starting cut of the side triangle completely. Under the same roof pressure conditions, the roof in the goaf is more likely to collapse.

[0017] Third, within the 10m range of the advanced area of the working face, prefabricated circumferential slots are made at different layers of the hard rock layer through the drilled boreholes, and hydraulic fracturing is performed at the slotting positions, thereby dividing the thick and hard roof into multiple thin and hard roofs. After the working face has passed, the roof is more likely to collapse, effectively reducing the cantilever roof length, not easily forming a large-area cantilever roof in the goaf, reducing the roof pressure, and improving the safety.

[0018] Fourth, the present invention breaks through the conventional method of hydraulic fracturing to cut the top of the key strata on one side of the roadway. The fracturing effect of this traditional construction method of hydraulic fracturing for roof cutting is generally average. It is proposed to use an abrasive water jet coal seam cutting and pressure relief device to prefabricate circumferential grooves in the hard thick rock strata at the target layer of the roadway, and conduct hydraulic fracturing at the grooving position. Through circumferential grooving, a large amount of high-pressure water is injected into the rock fractures, increasing the moisture content of the rock strata, reducing the strength of the rock strata. Through layered roof cutting, the thick and hard roof strata are divided into thin and hard rock strata, and the roof rock strata are more likely to collapse after the working face has passed. BRIEF DESCRIPTION OF THE DRAWINGS

[0019] Figure 1 In the figures: a is a schematic plan view of the end caving roof after the mining of the working face without implementing layered hydraulic fracturing; b is a schematic plan view of the end caving roof after the mining of the working face with implementing layered hydraulic fracturing; in the figures: 1 - borehole; 2 - hydraulic fracturing area; 3 - end caving roof; 4 - abrasive water jet combined with hydraulic fracturing boreholes in the advanced area. Figure 2 is Figure 1 Section A - A of b; Figure 3 is Figure 1 Section B - B of b. DETAILED DESCRIPTION OF THE INVENTION

[0020] In order to more clearly understand the above - mentioned objects, features, and advantages of the present invention, the present invention will be further described in detail below in conjunction with the drawings and specific embodiments. It should be noted that, without conflict, the embodiments of the present application and the features in the embodiments can be combined with each other.

[0021] In the following description, many specific details are set forth in order to fully understand the present invention. However, the present invention can also be implemented in other ways different from those described herein. Therefore, the protection scope of the present invention is not limited by the specific embodiments disclosed below.

[0022] A method for treating the end caving roof by combining abrasive water jet grooving with hydraulic fracturing for layered roof cutting. Use a drilling rig to drill boreholes into the hard roof of the target rock strata simultaneously on both sides of the roadway, namely the coal pillar side and the mining side. In the advanced range of the working face, use an abrasive water jet coal seam cutting and pressure relief device to cut circumferential grooves parallel to the roof plane at different layers of the target rock strata through the boreholes, and conduct hydraulic fracturing at the grooving position, dividing the thick and hard roof into multiple layers of thin and hard roofs.

[0023] Furthermore, it includes the following steps: S1, collect the engineering geological data of the roadway and the working face; S2, determine the borehole angle, borehole spacing, circumferential grooving, and hydraulic fracturing parameters; S3. Drill holes into the roof along the advancing direction of the working face according to the design parameters until the target horizon is reached, obtaining multiple hydropressurized holes arranged in parallel. S4. In the advanced area of the working face, use an abrasive water jet coal seam slotting and pressure relief device to carry out circumferential grooving in the drill hole and conduct hydraulic fracturing at the grooving position. S5. After the hydraulic fracturing is completed, use a peephole instrument to analyze the fracturing effect. Then, the working face starts to be mined, and specifically analyze the hanging roof area at the end of the working face, the gas concentration at the corner and the stress of the end support before and after roof cutting.

[0024] Furthermore, in step S1, the geological data includes: roadway driving operation regulations and support section drawings, mining and excavation engineering plans, surrounding borehole columnar diagrams, geological reports, and preliminary mine design specifications.

[0025] In step S2, determine the drill hole angle, drill hole spacing, target horizon thickness and required circumferential grooving position through theoretical calculation, on-site peeping and laboratory tests; specifically including the following steps: S2.1 Determine the drill hole spacing: The drill hole spacing is 2 times the crack propagation radius of hydraulic fracturing, specifically: D = 2 R ; In the formula: R — Crack propagation radius, m; D— Drill hole spacing, m; The crack propagation radius of hydraulic fracturing is based on the following formula: ; In the formula: Q — Injection rate of hydraulic fracturing fluid, L / min; E — Elastic modulus of rock, MPa; σ y — Rock fracture strength, MPa; K I — Rock crack propagation toughness; S2.2. Determine the drill hole angle: Before the working face is mined, according to the designed drill hole spacing, combined with the hanging roof occurrence at the end during the mining process of the working face under similar engineering geological conditions, use a drill rig to drill holes perpendicular to the roof on the side close to the coal pillar, and drill the holes to the midpoint position of the side triangular plate parallel to the starting cut line on the side close to the mining side. Specifically, there are the following formulas: ; ; In the formula: α— Drill hole azimuth angle, °; β — Drill hole inclination angle, °; D — Drill hole spacing, m; y — Length of the side triangular plate parallel to the starting cut, m; h — Vertical height from the fracturing horizon to the roof (m); Calculated according to the above formula: When the fracturing interval is located in conglomerate or limestone with uniaxial compressive strength greater than 60 MPa, the crack propagation toughness is taken as 1.2; the borehole azimuth angle α: ; the borehole dip angle β : ; When the fracturing interval is located in fine-grained sandstone or medium-grained sandstone with uniaxial compressive strength ranging from 50 to 60 MPa, the crack propagation toughness is taken as 1.0; the borehole azimuth angle α: ; the borehole dip angle β: ; When the fracturing interval is located in fine-grained sandstone or medium-grained sandstone with uniaxial compressive strength ranging from 40 to 50 MPa, the crack propagation toughness is taken as 0.8; the borehole azimuth angle α: ; the borehole dip angle β: ; S2.3. Determine the number and position of the holes according to the thickness of the hard rock layer at the target interval: When the thickness of the hard rock layer is less than 5 m, prefabricate 1 circumferential groove in the borehole and the groove is arranged at the 1 / 2 position of the rock layer; when the thickness of the cut hard rock layer is 5 - 10 m, prefabricate 2 circumferential grooves in the borehole and the grooves are arranged at the 1 / 3 position of the rock layer; when the thickness of the cut hard rock layer is 10 - 15 m, prefabricate 3 circumferential grooves in the borehole and each groove is successively arranged at the positions of 1 / 4, 1 / 2, and 3 / 4 from the boundary of the rock layer. S2.4. Obtain the required abrasive water jet flow rate, abrasive water jet pressure, abrasive concentration, and water injection pressure by testing the compressive and tensile strengths of the required fractured thick rock layer in the laboratory.

[0026] Furthermore, step S3 is to use a drilling rig to drill holes along the advancing direction of the working face towards the roof on both sides of the roadway, namely the coal pillar side and the mining side, according to the borehole design parameters in S2 until the target interval is reached.

[0027] Furthermore, step S4 includes the following steps: S4.1. Withdraw the drilling rig, insert the abrasive water jet coal seam slotting and pressure relief device into the drilled hole, rotate the drill pipe at a constant speed, and conduct circumferential grooving according to the design parameters in S2. S4.2. Start hydraulic fracturing using the high-pressure water pump of the abrasive water jet coal seam slotting and pressure relief device. Each fracturing time is ≥ 30 minutes, and the water pressure drops significantly, indicating that the fracturing is completed. Due to the high pressure, new cracks appear in the fractured rock layer and begin to expand. The water pressure drops slightly, and the cracks spread continuously in all directions. Then the water pressure drops significantly, indicating that the fracturing is completed.

[0028] In step S5: After the hydraulic fracturing is completed, use a borehole peephole to peep the borehole again to analyze the hydraulic fracturing effect; after the hydraulic fracturing is completed, the working face starts mining, and compare and analyze the suspended roof area at the end of the working face, the gas concentration at the corner, and the stress of the support before and after roof cutting.

[0029] The following further illustrates the specific embodiments of the present invention in conjunction with the accompanying drawings through specific examples. The accompanying drawings are schematic diagrams under the working conditions of Embodiment 1.

[0030] Embodiment 1: 10108 fully-mechanized coal mining face in a certain mine in Shanxi: A method for using abrasive water jet grooving combined with hydraulic fracturing for layered roof cutting to handle the end suspended roof. The specific method is carried out according to the following steps: Step I: Go to the site to collect engineering geological data of the roadway and the working face, such as: roadway tunneling operation regulations and support section drawings, mining engineering plans, surrounding borehole columnar diagrams, geological reports, mine preliminary design specifications, etc. According to the collected mine data, it can be known that the average thickness of the No. 10 coal seam is 6.2 m, and the full-height coal mining method with large mining height is adopted, with a mining height of 6.2 m. In addition, the 10108 track gateway is driven along the roof, with a length of 1274 m, a width of 5.5 m, and a height of 4.0 m. Its direct roof is limestone (9.9 m thick), sandy mudstone (6.4 m thick), mudstone (1.2 m thick), fine-grained sandstone (2.6 m thick), limestone (2.7 m thick), etc.

[0031] Step II: Use a peephole to conduct roof peeping at different positions in the roadway in sequence, record the roof lithology and thickness at different positions, determine the fracturing layer as the direct roof limestone, determine the injection rate of the hydraulic fracturing fluid as 50 L / min according to theoretical calculations and laboratory tests. The elastic modulus of the limestone is 10.0 GPa, and the average compressive strength is 60.1 MPa. According to the above calculation formula, the borehole spacing is calculated to be 3.1 m. According to the occurrence of the end suspended roof during the working face mining process under similar engineering geological conditions, the length of the lateral triangular plate parallel to the open-off cut is 7.8 m, and the calculated borehole azimuth angle is 51.6°, and the inclination angle is 63.4°. To ensure the fracturing effect and facilitate on-site construction, the borehole spacing is taken as 3.0 m, the azimuth angle is taken as 50.0°, and the inclination angle is 60.0°. Since the direct roof limestone is 9.9 m thick, the positions for prefabricating circumferential grooves in the borehole are calculated to be 3.3 m and 6.6 m.

[0032] Step III: Use drilling rigs to drill holes into the roof respectively at 0.5 m away from the mining side and the coal pillar side on both sides of the roadway until the limestone roof is penetrated. The required designed circumferential groove parameters need to be determined by comprehensively considering the thickness of the hard rock layer and the uniaxial compressive strength results. The specific circumferential groove parameters are as follows: the abrasive water jet pressure is 35 - 40 MPa, the nozzle diameter is φ 1.0 - 1.5 mm, the abrasive particle size is 60 - 80 mesh, and the abrasive concentration is 5% - 10%.

[0033] Step Ⅳ: Within a range of 10 m in the advance area of the working face, circumferential grooving is carried out at 3.3 m and 6.6 m respectively by using the abrasive water jet coal seam slotting pressure relief device through the pre-drilled boreholes. After the grooving is completed, the device is continuously used for hydraulic fracturing.

[0034] Step Ⅴ: After the circumferential hydraulic grooving of the borehole is completed, the borehole is peeped again by using a peephole to analyze the hydraulic fracturing effect. After the hydraulic fracturing is completed, the working face starts to be mined. During the mining, the working resistance of the support is monitored, and the cantilever roof area, corner gas concentration and the force on the end support at the working face before and after roof cutting are compared and analyzed. Through monitoring and calculation, it is obtained that after the layered hydraulic fracturing, the cantilever roof area at the end is reduced from 31.0 m 2 to 3.6 m 2 , the gas at the upper corner of the working face is reduced from 0.75% to about 0.3%, and the working resistance of the end support is reduced by up to 23.6%.

[0035] Example 2: A 30109 fully mechanized caving face in a certain No. 2 Coal Mine in Shanxi: A method for treating the cantilever roof at the end by using abrasive water jet grooving combined with hydraulic fracturing and layered roof cutting, and the method is carried out according to the following steps: Step Ⅰ: Go to the site to collect the engineering geological data of the roadway and the working face, such as: the roadway driving operation regulations and support section drawings, the mining engineering plan, the surrounding borehole columnar diagrams, the geological reports, the preliminary design specifications of the mine, etc. According to the collected mine data, it is known that the average thickness of the No. 3 coal seam is 6.1 m, the longwall fully mechanized top coal caving mining method is adopted, the mining height is 3.3 m, the mining and caving ratio is 1:0.85, and the gob management adopts the full caving method. In addition, the 30109 track gateway is driven along the bottom, with a length of 2278 m, a width of 5.0 m and a height of 3.5 m. Its direct roof is fine sandstone (15.0 m thick), mudstone (2.3 m thick), siltstone (4.5 m thick), mudstone (1.7 m thick), fine sandstone (1.8 m thick), etc.

[0036] Step Ⅱ: Use a peephole to peep the roof at different positions in the roadway in turn, record the roof lithology and thickness at different positions, determine the fracturing layer as the direct roof fine sandstone, determine the injection rate of the hydraulic fracturing fluid as 80 L / min according to theoretical calculation and laboratory tests, the elastic modulus of the fine sandstone is 5.3 GPa, and the average compressive strength is 68.6 MPa. The borehole spacing is calculated to be 3.4 m. According to the occurrence of the cantilever roof at the end during the mining process of the working face under similar engineering geological conditions, the length of the lateral triangular plate parallel to the starting cut is 8.5 m, and the borehole azimuth angle is calculated to be 51.3°, and the inclination angle is 59.1°. To ensure the fracturing effect and facilitate on-site construction, the borehole spacing is taken as 3.0 m, the azimuth angle is taken as 50.0°, and the inclination angle is 60.0°. Since the direct roof fine sandstone is 15.0 m thick, the positions for prefabricating circumferential grooves in the borehole are calculated to be 3.8 m, 7.6 m, and 11.4 m.

[0037] Step Ⅲ: Drill holes into the roof respectively on both sides of the roadway at a distance of 0.5 m from the mining side and the coal pillar side until the thick and hard strata of the immediate roof are penetrated. The required designed circumferential slotting parameters need to be determined by comprehensively considering the thickness of the hard strata and the results of uniaxial compressive strength. The specific circumferential slotting parameters are as follows: the abrasive water jet pressure is 40 - 50 MPa, the nozzle diameter φ is 1.0 - 1.5 mm, the abrasive particle size is 55 - 85 mesh, and the abrasive concentration is 7% - 12%.

[0038] Step Ⅳ: Within a range of 10 m in the advanced area of the working face, use the abrasive water jet coal seam slotting pressure relief device through the pre-drilled holes to conduct circumferential slotting at 3.8 m, 7.6 m, and 11.4 m respectively. After the slotting is completed, continue to use this device for hydraulic fracturing. The trajectory of crack propagation is a circular crack parallel to the roof bedding plane; Step Ⅴ: After the circumferential hydraulic slotting of the borehole is completed, use the borehole peephole to peep the borehole again to analyze the effect of hydraulic fracturing; after the hydraulic fracturing is completed, the working face starts to be mined. During the mining process, monitor the working resistance of the support, compare and analyze the cantilever roof area at the working face end, the gas concentration at the corner, and the force on the end support. Through monitoring and calculation, it is obtained that after the layered hydraulic fracturing, the cantilever roof area at the end is reduced from 36.2 m 2 to 4.1 m 2 , the gas at the upper corner of the working face is reduced from 0.7% to about 0.2%, and the working resistance of the end support is reduced by up to 21.3%.

[0039] Finally, it should be noted that the above embodiments are only used to illustrate the technical solutions of the present invention, rather than to limit them; although the present invention has been described in detail with reference to the foregoing embodiments, those of ordinary skill in the art should understand that they can still modify the technical solutions recorded in the foregoing embodiments, or perform equivalent replacements for some or all of the technical features; and these modifications or replacements do not make the essence of the corresponding technical solutions deviate from the scope of the technical solutions of the embodiments of the present invention.

Claims

1. A method for treating the overhanging roof at the end by using a frosted water jet cutting groove in combination with hydraulic fracturing for layered roof cutting, characterized in that: Using a drill rig, drill holes into the hard roof of the target rock formation simultaneously on both sides of the roadway, namely the coal pillar side and the mining side. In the area ahead of the working face, use an abrasive water jet coal seam cutting and pressure relief device to cut circumferential grooves parallel to the roof plane at different levels of the target rock formation through the drill holes, and conduct hydraulic fracturing at the grooving positions to divide the thick and hard roof into multiple thin and hard roofs.

2. The method according to claim 1, wherein: It includes the following steps: S1. Collect the engineering geological data of the roadway and the working face; S2. Determine the drilling angle, drilling spacing, circumferential grooving and hydraulic fracturing parameters; S3. According to the design parameters, drill holes into the roof along the advancing direction of the working face until the target layer is reached, obtaining multiple hydraulically relieved holes arranged in parallel; S4. In the area ahead of the working face, use an abrasive water jet coal seam cutting and pressure relief device to conduct circumferential grooving in the drill holes and carry out hydraulic fracturing at the grooving positions; S5. After the hydraulic fracturing is completed, use a peephole to analyze the fracturing effect, and then the working face starts mining, and specifically analyze the hanging roof area at the working face end, the gas concentration at the corner and the stress of the end support before and after roof cutting.

3. The method according to claim 2, wherein: In step S1, the geological data includes: the roadway tunneling operation regulations and support section drawings, the mining engineering plan, the surrounding drill hole columnar diagrams, the geological reports, and the preliminary design specifications of the mine.

4. The method according to claim 2, wherein: In step S2, through theoretical calculations, on-site peeping and laboratory tests, determine the drilling angle, drilling spacing, thickness of the target layer and the required circumferential grooving positions; specifically, it includes the following steps: S2.1 Determine the drilling spacing: The drilling spacing is 2 times the crack propagation radius of hydraulic fracturing, specifically: D = 2 R ; Wherein: R — Crack propagation radius, m; D— Borehole spacing, m; The propagation radius of hydraulic fracturing cracks is based on the following formula: ; where: Q — Injection rate of hydraulic fracturing fluid, L / min; E — Elastic modulus of rock, MPa; σ y — Fracture strength of rock, MPa; K I — Crack propagation toughness of rock; S2.2 Determine the drilling angle: Before the working face is mined, according to the designed hole spacing and in combination with the occurrence of the end - hanging roof during the mining process of the working face under similar engineering geological conditions, use a drill rig to drill holes perpendicular to the roof on the side close to the coal pillar, and drill the holes to the mid - point position of the side - triangle plate parallel to the starting cut side line on the side close to the mining side. The specific formula is as follows: ; ; Where: α—the azimuth angle of the drill hole, °; β —the dip angle of the drill hole, °; D —the hole spacing, m; y — Length of the side triangular plate parallel to the starting cut roadway, m; h —Vertical height of the fracturing interval from the roof, m; Calculated according to the above formula: When the fracturing interval is located in conglomerate or limestone with a uniaxial compressive strength greater than 60 MPa, the crack propagation toughness is taken as 1.2; the borehole azimuth α: ; the borehole inclination β : ; When the fracturing interval is located in fine-grained sandstone or medium-grained sandstone, with the uniaxial compressive strength ranging from 50 to 60 MPa, the crack propagation toughness is taken as 1.0; the borehole azimuth angle α: ; the borehole dip angle β: ; When the fracturing interval is located in fine-grained sandstone or medium-grained sandstone, with the uniaxial compressive strength ranging from 40 to 50 MPa, the crack propagation toughness is taken as 0.8; borehole azimuth α: ; borehole dip angle β: ; S2.

3. Determine the number and positions of the holes to be drilled according to the thickness of the hard rock formation in the target layer: When the thickness of the hard rock formation is less than 5m, prefabricate 1 circumferential groove in the drill hole and the groove is set at the 1 / 2 position of the rock formation; when the thickness of the cut hard rock formation is 5 - 10m, prefabricate 2 circumferential grooves in the drill hole and the grooves are set at the 1 / 3 position of the rock formation; when the thickness of the cut hard rock formation is 10 - 15m, prefabricate 3 circumferential grooves in the drill hole and each groove is sequentially set at the 1 / 4, 1 / 2, 3 / 4 positions from the boundary of the rock formation; S2.

4. Through laboratory tests on the compressive and tensile strengths of the thick rock formation to be fractured, obtain the required abrasive water jet flow rate, abrasive water jet pressure, abrasive concentration and water injection pressure.

5. The method according to claim 2, wherein: Step S3 is to drill holes into the roof along the advancing direction of the working face on both sides of the roadway, namely the coal pillar side and the mining side, according to the drilling design parameters in S2 until the target layer is reached.

6. The method according to claim 2, characterized in that: Step S4 includes the following steps: S4.

1. Withdraw the drill rig, insert the abrasive water jet coal seam cutting and pressure relief device into the drilled hole, rotate the drill pipe at a constant speed, and conduct circumferential grooving according to the design parameters in S2; S4.

2. Start hydraulic fracturing using the high-pressure water pump of the abrasive water jet coal seam cutting and pressure relief device, with each fracturing time ≥ 30 minutes and a significant reduction in water pressure, at which point the fracturing is completed.

7. The method according to claim 2, wherein: In step S5: After the hydraulic fracturing is completed, the borehole is peeped again using a peephole instrument to analyze the hydraulic fracturing effect; after the hydraulic fracturing is completed, the working face starts to be mined, and the hanging roof area at the end of the working face, the gas concentration at the corner, and the stress on the support are compared and analyzed before and after roof cutting.

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