Roof fracturing method and stop mining isobaric method for coal mining face in final mining stage

Abstract

The application discloses a roof fracturing method for a coal mine working face in a final mining stage and a stop-mining isobaric method. The roof fracturing method for the coal mine working face in the final mining stage comprises the following steps: S1, calculating the positions of fracture lines L1 and L2, the distance between the fracture line L1 and a withdrawing roadway being W, and the distance between the fracture line L2 and the withdrawing roadway being W'; S2, determining the positions of high-positioned roof and low-positioned roof which play a main control role in roof caving, the height of the high-positioned roof from the working face being greater than the height of the low-positioned roof from the working face; and S3, fracturing the low-positioned roof and part of the high-positioned roof between the fracture lines L1 and L2, and fracturing the low-positioned roof and the high-positioned roof on the rear side of the fracture line L2. In the roof fracturing method for the coal mine working face in the final mining stage, the application has pertinence to different mines compared with a traditional engineering analogy method, good pressure-relief effect is achieved, and underground equipment is easy to withdraw after the coal body is mined.

Description

Technical Field

[0001] This invention relates to the field of coal mine safety technology, specifically to a method for inducing roof fracturing and a method for stopping mining and equalizing pressure in the final stage of coal mine working face. Background Technology

[0002] Coal mining typically employs a longwall mining method, which involves starting mining from a distance away from the working face and gradually moving towards the retreat roadway. In the final stages of mining, it is generally necessary to fracturate the hard, thick roof covering the coal seam to relieve pressure and control roof collapse, thereby ensuring safe production.

[0003] In existing technologies, the relevant parameters for roof fracturing treatment are generally determined through engineering analogy. For example, the decompression scheme of one mine is directly applied to another mine, or the decompression scheme of the entire mining stage is directly applied to the roof of the final mining stage. This method often lacks a reasonable basis for designing fracturing parameters, which can easily lead to poor decompression effects in coal mines and fail to eliminate safety hazards caused by decompression problems. Summary of the Invention

[0004] This invention is based on the inventor's discoveries and understanding of the following facts and problems:

[0005] The inventors' research revealed that, in the final stages of coal mining, the roof cracking presents the following two problems.

[0006] 1. If the cracked area of ​​the roof is too far from the retreat roadway, the length of the roof suspended will be large, which will lead to greater underground pressure in the coal mine and make it prone to dynamic disasters.

[0007] 2. If the roof fracturing zone is too close to the retreat roadway, although a good pressure relief effect can be achieved, the rock falling from above the working face will accumulate on the underground equipment, making it difficult to withdraw the equipment. Therefore, it is necessary to perform zoned fracturing of the roof in the final mining stage to achieve a better pressure relief effect and facilitate the withdrawal of underground equipment.

[0008] The present invention aims to at least partially solve one of the technical problems in the related art.

[0009] Therefore, this invention proposes a method for fracturing the roof of a coal mine working face in the final stage of mining. This method calculates the location of the fracture line and performs zoned fracturing treatment on the roof according to the fracture line, which can achieve better pressure relief, reduce dynamic disasters, and facilitate the removal of underground equipment.

[0010] This invention also proposes a method for stopping mining under equal pressure.

[0011] The method for inducing roof fracturing in the final stage of coal mine working face according to an embodiment of the present invention includes the following steps:

[0012] S1: Calculate the positions of fracture lines L1 and L2, where the distance between fracture line L1 and the retreat roadway is... W The distance between the fracture line L2 and the retreat roadway ,

[0013]

[0014]

[0015] in, W 1 represents the width of the retreat tunnel, in meters (m). W 2 represents the distance between the support frame and the stack, in meters. W 3 represents the control distance of the working face support, in meters; D The width of the remaining coal pillar, in meters;

[0016] S2: Determine the positions of the high-level roof plate and the low-level roof plate that play a major role in controlling the roof collapse, wherein the height of the high-level roof plate from the working surface is greater than the height of the low-level roof plate from the working surface;

[0017] S3: Perform cracking treatment on the low-position top plate and part of the high-position top plate located between the fracture lines L1 and L2, and perform cracking treatment on the low-position top plate and the high-position top plate located behind the fracture line L2.

[0018] In the roof fracturing method of the coal mine working face in the final mining stage of this invention, the positions of fracture lines L1 and L2 are calculated using parameters such as the width of the coal mine's retreat roadway and the distance between supports and stackers. This allows for a relatively accurate determination of the areas in the roof that require fracturing treatment in the final mining stage. On the one hand, this avoids the fracturing area being too far from the retreat roadway, resulting in excessive roof overhang and high underground pressure, which could easily lead to dynamic disasters. On the other hand, it also avoids the fracturing area being too close to the retreat roadway, causing the roof to collapse and accumulate on underground equipment, making it difficult to evacuate the equipment.

[0019] Identifying the locations of the high-level and low-level roof sections that play a major role in controlling roof collapse, and then fracturing them accordingly, is more effective and can achieve better pressure relief and hazard mitigation results. Compared to traditional engineering analogy methods, the roof fracturing method for the final mining stage of this invention is more targeted to different mines, achieving better pressure relief while facilitating the removal of underground equipment after coal mining is completed.

[0020] In some embodiments, in step S1, the remaining coal pillar width D Calculated using the following formula:

[0021]

[0022]

[0023] =

[0024]

[0025] in, The average stress of the remaining coal pillar is given in kPa. The total load of the remaining coal pillar is kN; The critical width, in meters, required to maintain the stability of the remaining coal pillar. The sloping length is taken as 1m; The strength of the coal pillar is given in kPa. The uniaxial compressive strength of the cubic coal sample is given in kPa. The depth is measured in meters (m).

[0026] In some embodiments, step S3 includes the following steps:

[0027] S31: Divide the roof above the coal seam in the retreat roadway into region A.

[0028] The roof above the coal seam in the return air roadway, located between the fracture lines L1 and L2, is designated as region B1.

[0029] The roof above the coal seam in the conveyor roadway, located between the fracture lines L1 and L2, is designated as region B2.

[0030] The roof above the coal seam of the return air roadway, located behind the fracture line L2, is divided into region C1.

[0031] The roof above the coal seam in the conveyor roadway, located behind the fracture line L2, is designated as region C2.

[0032] S32: Within region A, boreholes are drilled from the retreat roadway to the roof above the coal face, such that the fracturing section of a portion of the borehole reaches the high roof behind the fracture line L1, and the fracturing section of a portion of the borehole reaches the low roof behind the fracture line L1.

[0033] And / or, within regions B1 and B2, boreholes are drilled from the return air roadway and the conveyor roadway toward the roof above the coal seam of the working face, such that the fracturing zone of the boreholes reaches the lower roof.

[0034] And / or, within the regions C1 and C2, boreholes are drilled from the return air roadway and the conveyor roadway toward the roof above the coal body of the working face, so that the fracturing section of the borehole reaches the low roof and the high roof.

[0035] S33: Crack the high-level top plate and / or the low-level top plate by drilling holes in the regions A, B1, B2, C1, and C2.

[0036] In some embodiments, in step S32, the boreholes in region A include high-level deep holes and low-level shallow holes, and the high-level deep holes and low-level shallow holes are arranged alternately along the extension direction of the retreat roadway. The fracturing section of the high-level deep holes reaches the high-level roof, and the fracturing section of the low-level shallow holes reaches the low-level roof.

[0037] In some embodiments, in step S32, the drill holes in region C1 are arranged at intervals along the return air roadway, and the drill holes in region C2 are arranged at intervals along the extension direction of the adhesive conveying roadway, and the number of drill holes in regions C1 and C2 is 2-6.

[0038] In some embodiments, in step S22, the drilling spacing d Calculated using the following formula: ,in, r Let be the crack radius, in meters.

[0039] In some embodiments, the method for fracturing the high-level top plate and / or the low-level top plate in step S33 includes at least one of the following: placing a fracturing agent in the fracturing section of the borehole and sealing the borehole, inducing the fracturing agent to react and fracturing the high-level top plate and / or the low-level top plate; sealing the fracturing section of the borehole and continuously injecting high-pressure fluid into the fracturing section to fracture the high-level top plate and / or the low-level top plate.

[0040] In some embodiments, in step S33, high-pressure fluid is used to cause cracks in the high-level top plate and / or the low-level top plate, and step S33 includes the following steps:

[0041] S331: Divide the fracture-causing section of the borehole into multiple borehole segments along the borehole extension direction;

[0042] S332: Seal both ends of each drilled section in sequence and continuously inject high-pressure fluid so that the high-level top plate and / or the low-level top plate can be layered along the drilled section and collapse in sequence.

[0043] In some embodiments, step S2, determining the positions of the high-level top plate and the low-level top plate, includes the following steps:

[0044] S21: Drill holes above the coal seam of the working face to observe and take rock cores to determine the physical and mechanical parameters of each layer of the roof;

[0045] S22: Based on the physical and mechanical parameters of each top plate, perform theoretical calculations to obtain the position data of the high-level top plate and the low-level top plate;

[0046] S23: Collect spatial location data of micro-seismic events that occur during the mining of the working face, and analyze the location data of the high-level roof and the low-level roof to obtain the second set of location data;

[0047] S24: Compare the position data one and the position data two. When the deviation between the position data one and the position data two meets the set standard, determine the position of the high-position top plate and the low-position top plate.

[0048] The method for stopping mining and equalizing pressure according to embodiments of the present invention includes the following steps:

[0049] Z1: The roof of the coal mine working face in the final mining stage is subjected to fracturing treatment according to the roof fracturing method in any of the above embodiments.

[0050] Z2: Conduct the final mining operation at the working face. After mining passes through the fracture line L2, stop mining and wait for the roof behind the fracture line L2 to collapse.

[0051] Z3: Continue mining the remaining coal seam and pass through fracture line L1;

[0052] Z4: Remove downhole equipment from the working face. Attached Figure Description

[0053] Figure 1 This is a schematic diagram of the location of fracture line L1 in the method for causing cracking of the roof of a coal mine working face in the final mining stage according to an embodiment of the present invention.

[0054] Figure 2 This is a schematic diagram of the cracking zone division of the roof cracking method in the final mining stage of the coal mine working face according to an embodiment of the present invention.

[0055] Figure 3 This is a schematic diagram of the drilling layout in the cracking zone of the roof cracking method for the coal mine working face in the final mining stage according to an embodiment of the present invention.

[0056] Figure 4 This is a cross-sectional view of the borehole layout in region A of the coal mine working face roof cracking method in the final mining stage of this embodiment of the invention.

[0057] Figure 5 This is a cross-sectional view of the borehole layout in areas B1 and B2 of the coal mine working face roof cracking method in the final mining stage of this invention.

[0058] Figure 6 This is a cross-sectional view of the borehole layout in areas C1 and C2 of the coal mine working face roof cracking method in the final mining stage of this embodiment of the invention.

[0059] Figure 7 This is a schematic diagram of the fracturing borehole arrangement in Embodiment 1 of the present invention.

[0060] Figure 8 This is a cross-sectional view of the arrangement of boreholes a and b in Embodiment 1 of the present invention.

[0061] Figure 9 This is a cross-sectional view of the arrangement of borehole c in Embodiment 1 of the present invention.

[0062] Figure 10 This is a cross-sectional view of the arrangement of borehole c in Embodiment 1 of the present invention.

[0063] Figure 11 This is a stress data diagram of the shallow hole measuring point in the return air channel of Embodiment 1 of the present invention.

[0064] Figure 12 This is a stress data diagram of the deep hole measuring point in the return air channel of Embodiment 1 of the present invention.

[0065] Figure 13 This is a schematic diagram of the location of the high-energy microseismic event in the fracturing section on January 23, according to Embodiment 1 of the present invention.

[0066] Figure 14 This is a schematic diagram of the location of the high-energy microseismic event in the fracturing section on January 24th, according to Embodiment 1 of the present invention.

[0067] Figure 15 This is a schematic diagram of the location of the high-energy microseismic event in the fracturing section on January 27th, according to Embodiment 1 of the present invention.

[0068] Figure 16 This refers to the statistical data on the step distance of the working surface in Embodiment 1 of the present invention.

[0069] Figure 17 This is a schematic diagram of roof fracture calculation D in the roof fracture method of the coal mine working face in the final mining stage of the present invention.

[0070] Figure 18 This is a mechanical model diagram for calculating D in the method for causing roof cracking in the coal mine working face during the final mining stage of this invention.

[0071] Figure label:

[0072] 1. Coal body; 2. Roof; 21. High roof; 22. Low roof; 3. Retreat roadway; 4. Return airway; 5. Glue transport roadway; 6. Working face. Detailed Implementation

[0073] Embodiments of the present invention are described in detail below, with examples of the embodiments illustrated in the accompanying drawings. The embodiments described below with reference to the accompanying drawings are exemplary and intended to explain the present invention, and should not be construed as limiting the present invention.

[0074] like Figures 1-6 As shown, the method for fracturing the roof of a coal mine working face in the final mining stage (hereinafter referred to as the method) according to an embodiment of the present invention includes the following steps:

[0075] S1: Calculate the positions of fracture lines L1 and L2. The distance between fracture line L1 and the retreat roadway 3 is... W The distance between the fracture line L2 and the retreat roadway 3 ,

[0076]

[0077]

[0078] in, W 1 represents the width of the retreat tunnel, in meters (m). W 2 represents the distance between the support frame and the stack, in meters. W 3 represents the distance between the 6 supports controlling the top of the working face, in meters. D The width of the remaining coal pillar is in meters.

[0079] Specifically, such as Figures 1-3 As shown, the black portion represents coal body 1, and the fracture line L1 is the last fracture line that caused the roof 2 to fracture during the final mining stage of working face 6. W The calculation measures the distance from the non-production side of the retreat roadway to fault line L1. Fault line L2 is the penultimate fault line of the roof that fractured during the final mining stage of working face 6. The calculation is based on the distance between the non-production side of the retreat roadway and the fracture line L2.

[0080] Existing roof fracturing methods generally do not consider the distance between the fracturing section and the retreat roadway 3. When the fracturing section is too far from the retreat roadway 3, the length of the suspended roof is too long, resulting in significant underground mine pressure and a high risk of dynamic disasters. Conversely, when the fracturing section is close to the retreat roadway 3, the collapsed roof and other debris accumulate on the underground equipment, making it difficult to evacuate. In the method of this embodiment, the location of the fracture line L1 is determined by calculation based on data such as the width of the retreat roadway and the distance between the supports and the stackers in the working face 6. Fracturing the roof at location L1 achieves both better pressure relief and facilitates the evacuation of underground equipment.

[0081] When mining face 6, mining can be suspended for a period of time after passing through fracture line L2. Since the collapse of the roof has a lag effect, the purpose of suspending mining is to wait for the roof behind fracture line L2 to bend, break, and fully collapse, thereby achieving a better pressure relief and hazard mitigation effect. Existing fracturing methods do not consider roof fracturing during the suspension phase. However, in this method, the location of fracture line L2 is determined through calculation, and fracturing treatment of the roof at L2 location can achieve a better pressure relief effect and reduce the occurrence of dynamic disasters.

[0082] S2: Determine the positions of the high-level roof slab 21 and the low-level roof slab 22, which play a major role in controlling the collapse of the roof slab. The height of the high-level roof slab 21 from the working surface 6 is greater than the height of the low-level roof slab 22 from the working surface 6.

[0083] The overlying rock of working face 6 is generally layered, and certain layers play a major role in controlling the collapse of the roof, namely the aforementioned high-level roof 21 and low-level roof 22. Roof collapse can be controlled by fracturing the high-level roof 21 and low-level roof 22; this fracturing treatment is more effective.

[0084] Specifically, core samples can be extracted by drilling into the rock strata, and then the lithology, thickness, and mechanical properties of the core samples can be analyzed to determine the stratigraphic positions of the high-level top plate 21 and the low-level top plate 22 that play a dominant role in the rock strata.

[0085] S3: Perform cracking treatment on the low-level roof plate and part of the high-level roof plate located between fracture lines L1 and L2, and perform cracking treatment on the low-level roof plate and the high-level roof plate located behind fracture line L2.

[0086] After the high-level and low-level roofs behind fracture line L2 are fractured, both the high-level roof and the low-level roof below the high-level roof 21 will gradually collapse, resulting in sufficient roof collapse and good pressure relief. Specifically, between fracture lines L1 and L2, in the area near the production side of the retreat roadway (behind fracture line L1), the low-level roof can be completely fractured, while the high-level roof above it can only be partially fractured. For example, only half the thickness of the high-level roof can be fractured, leaving half its thickness intact. The retained high-level roof provides temporary support to the overlying strata, reducing pressure on downhole equipment and facilitating its removal.

[0087] In other embodiments, in the area near the production side of the retreat roadway (behind fracture line L1), the lower roof can be completely fractured, while only two-thirds of the upper roof thickness is treated. It is understood that the thickness of the upper roof fracture treatment may vary in different mines, but a portion of the upper roof should be preserved for temporary support.

[0088] In the roof fracturing method of the coal mine working face in the final mining stage of this invention, the positions of fracture lines L1 and L2 are calculated using parameters such as the width of the coal mine's retreat roadway and the distance between supports and stackers. This allows for a relatively accurate determination of the areas in the roof that require fracturing treatment in the final mining stage. On the one hand, this avoids the fracturing area being too far from the retreat roadway, resulting in excessive roof overhang and high underground pressure, which could easily lead to dynamic disasters. On the other hand, it also avoids the fracturing area being too close to the retreat roadway, causing the roof to collapse and accumulate on underground equipment, making it difficult to evacuate the equipment.

[0089] Identifying the locations of the high-level and low-level roof sections that play a major role in controlling roof collapse, and then fracturing them accordingly, is more effective and can achieve better pressure relief and hazard mitigation results. Compared to traditional engineering analogy methods, the roof fracturing method for the final mining stage of this invention is more targeted to the mine, achieving better pressure relief while facilitating the removal of underground equipment after coal mining is completed.

[0090] In some embodiments, in step S1, the width of the remaining coal pillar D Calculated using the following formula

[0091]

[0092]

[0093] =

[0094]

[0095] in, The average stress of the remaining coal pillar is given in kPa. The total load of the remaining coal pillar is kN; The critical width, in meters, required to maintain the stability of the remaining coal pillar. The sloping length is taken as 1m; The strength of the coal pillar is given in kPa. The uniaxial compressive strength of the cubic coal sample is given in kPa. The depth is measured in meters (m).

[0096] Specifically, such as Figure 17 , Figure 18 As shown, a mechanical model for the stability analysis of the remaining coal pillar can be established. Within a certain range in front of the fracture location, the roof is subjected to the self-weight of the overlying strata and the transferred load from the goaf of the working face, forming a stress-increased zone. In this zone, the roof is subjected to a concentration factor of... K Advanced support pressure q 1( x Its function is to bear the weight of the upper caving strata within a certain range behind the fault location. q 2. Function.

[0097] In the final mining stage, the remaining coal pillar is affected not only by the overlying strata, but also by the load transferred from the roof of the retreat roadway and the working face roof. P 2, P 3. Function. The roof of the retreat roadway can be considered as a beam with fixed supports at both ends, from which we can obtain:

[0098]

[0099] in, q 1( x () represents the pre-fracture bearing pressure, in kPa; t The sloping length is taken as 1m; f 1 represents the working resistance of the stack frame, in kPa; w 1 represents the width of the retreat tunnel, in meters (m).

[0100] If the working face roof is considered as a cantilever beam and fixed to the remaining coal pillars, then:

[0101]

[0102] in, q 2 represents the bearing pressure after fracture, in kPa; f 2 represents the working resistance of the working face support, in kPa; w 2 represents the top distance of the support frame, in meters.

[0103] Based on the mineral pressure manifestation pattern of the overlying strata, the distribution of the pre-support stress is as follows:

[0104]

[0105] in, K The stress concentration factor; H The thickness of the overlying rock strata is in meters (m). γ The average unit weight of the overlying rock strata is kN / m³. 3 ; f The coefficient of friction; h The mining height is in meters (m). λ The lateral pressure coefficient, θ Take 1 / λ .

[0106] Force exerted by the broken top plate on the direct jacking q 2. Support force of the retraction tunnel support frame f 1. Working resistance of the working face support f If 2 can be considered as a uniformly distributed force, then the remaining coal pillar is subjected to the load of the overlying strata:

[0107]

[0108] The remaining coal pillar is subjected to loads transferred from the retreat roadway and loads from the overlying strata.

[0109]

[0110] The remaining coal pillar is subject to loads transmitted from the working face:

[0111]

[0112] The total load of the remaining coal pillars at the working face is:

[0113]

[0114] In the final mining stage, fracture zones appear on both sides of the remaining coal pillar. According to the Mohr-Coulomb failure criterion, the width of the fracture zone can be calculated:

[0115]

[0116] in, C The coal body cohesion, kPa; φ The friction angle within the coal is denoted as .

[0117] Because the fractured zone is severely damaged, its contribution can be ignored when calculating the bearing capacity of the coal pillar. Therefore, the average stress on the remaining coal pillar in the working face is:

[0118]

[0119] Calculate the coal pillar strength using the Bieniaski formula:

[0120]

[0121] in, σ p The strength of the coal pillar is given in kPa. σ 1 represents the uniaxial compressive strength of a cubic coal sample.

[0122] Finally, let σ a = σ p Solve for the critical width at which the remaining coal pillar maintains stability. w 0. D Greater than or equal to w 0 is sufficient.

[0123] In some embodiments, step S3 includes the following steps:

[0124] S31: The roof above the coal seam in the production side of the retreat roadway is designated as area A.

[0125] The roof above the coal seam in the return air roadway, located between fracture lines L1 and L2, is designated as region B1.

[0126] The roof above the coal seam in the production rib of the conveyor roadway, located between fracture lines L1 and L2, is designated as region B2.

[0127] The roof above the coal seam of the return air roadway, located behind the fracture line L2, is designated as region C1.

[0128] The roof above the coal seam of the production side of the conveyor roadway, which is located behind the fracture line L2, is divided into region C2;

[0129] S32: Within area A, boreholes are drilled from the retreat roadway 3 to the roof above the coal seam of working face 6, such that the fracturing zone of some boreholes reaches the high roof 21 behind fracture line L1, and the fracturing zone of some boreholes reaches the low roof 22 behind fracture line L1.

[0130] And / or, within regions B1 and B2, boreholes are drilled from the return air roadway 4 and the conveyor roadway 5 toward the roof above the coal seam of the working face 6, such that the fracturing zone of the boreholes reaches the lower roof 22.

[0131] And / or, within regions C1 and C2, boreholes are drilled from the return air roadway 4 and the conveyor roadway 5 toward the roof above the coal seam of the working face 6, so that the fracturing section of the boreholes reaches the low roof 22 and the high roof 21.

[0132] S33: Crack the high-level top plate and / or low-level top plate by drilling in areas A, B1, B2, C1, and C2.

[0133] like Figures 1-6 As shown, with the direction of advancement of working face 6 as the front, return air roadway 4 and conveyor roadway 5 are located on the left and right sides of working face 6, respectively, and the retreat roadway is located in front of working face 6. Areas A, B1, B2, C1, and C2 are all construction areas.

[0134] like Figure 4 As shown, area A is located on the side of the retreat roadway 3 near the working face 6. Construction workers can drill holes within the retreat roadway 3. The starting point of the borehole is within the retreat roadway 3, and the fracturing section of the borehole is located behind the fracture line L1. The entire borehole is located within area A. Simultaneously, the borehole inclination angle can be different, so that the fracturing section of some boreholes reaches the high roof 21 behind the fracture line L1, and the fracturing section of some boreholes reaches the low roof 22 behind the fracture line L1.

[0135] like Figure 5 As shown, area B1 is located on the side of the return air roadway 4 near the working face 6. Construction workers can drill holes in the return air roadway 4. The starting point of the hole is in the return air roadway 4. The cracking section of the hole is located between the fracture lines L1 and L2 and reaches the low top plate 22. The entire hole is located in area B1.

[0136] Area B2 is located on the side of the haulage roadway 5 near the working face 6. Construction workers can drill holes in the haulage roadway 5. The starting point of the drilling is in the haulage roadway 5. The cracking section of the drilling is located between the fracture lines L1 and L2 and reaches the low top plate 22. The entire drilling is located in area B2.

[0137] like Figure 6As shown, area C1 is located on the side of the return air roadway 4 near the working face 6. Construction workers can drill holes in the return air roadway 4. The starting point of the hole is in the return air roadway 4. The cracking section of the hole is located behind the fracture line L2 and reaches the high-level roof 21 and the low-level roof 22. The entire hole is located in area C1.

[0138] Area C2 is located on the side of the haulage roadway 5 near the working face 6. Construction workers can drill holes in the haulage roadway 5. The starting point of the drilling is in the haulage roadway 5. The cracking section of the drilling is located behind the fracture line L2 and reaches the high top plate 21 and the low top plate 22. The entire drilling is located in area C2.

[0139] Dividing the area of ​​the top plate that needs to be cracked facilitates construction. After the drilling is completed, the cracking section of the drilled hole can be used to carry out subsequent cracking treatment on the above-mentioned high-level top plate 21 and low-level top plate 22.

[0140] In some embodiments, in step S32, the boreholes in region A include high-level deep holes and low-level shallow holes, and the high-level deep holes and low-level shallow holes are arranged alternately along the extension direction of the retreat roadway. The fracturing section of the high-level deep holes reaches the high-level roof, and the fracturing section of the low-level shallow holes reaches the low-level roof.

[0141] like Figure 3 , Figure 4 As shown, in area A, high-level deep boreholes and low-level deep boreholes are arranged alternately. The purpose of this arrangement is to simultaneously fracturing the high-level roof 21 and low-level roof 22 located behind the fault line L1. Compared to treating only the high-level roof 21, this allows both the high-level roof 21 and low-level roof 22 behind the fault line L1 to collapse more quickly and completely. After the coal body has been mined past the fault line L1, the remaining coal body can be quickly mined out, and then the underground equipment can be withdrawn. Because both the high-level roof 21 and low-level roof 22 behind the fault line L1 are fractured, the roof behind the fault line L1 will collapse more quickly and completely, preventing the roof behind the fault line L1 from remaining suspended and thus damaging the coal body on the other side of the withdrawal roadway 3.

[0142] Furthermore, it should be understood that within area A, only a portion of the high-level roof 21 can be fractured. Specifically, the fractured section of the high-level deep borehole can cover only half the thickness of the high-level roof 21. The uncovered portion of the high-level roof 21 in the fractured section provides temporary support to the overlying strata. This treatment achieves better pressure relief and reduces the pressure on underground equipment, facilitating the removal of underground equipment after coal mining is completed.

[0143] In some embodiments, in step S32, such as Figure 3As shown, in step S32, the drill holes in region C1 are arranged at intervals along the return air roadway, and the drill holes in region C2 are arranged at intervals along the extension direction of the adhesive conveying roadway, and the number of drill holes in regions C1 and C2 is 2-6.

[0144] Specifically, the number of boreholes in regions C1 and C2 can be set to four, and the distance between any two adjacent boreholes is twice the cracking radius. This setting allows the top plate cracks between adjacent boreholes to be connected, and the length of the cracking treatment of the top plate in regions C1 and C2 reaches six times the cracking radius, thereby achieving a better pressure relief effect.

[0145] In other embodiments, the number of boreholes in regions C1 and C2 is two, three, five, six, etc.

[0146] In some embodiments, in step S22, the drilling spacing d Calculated using the following formula: ,in, r Let be the crack initiation radius, in meters (m). The purpose of setting the borehole spacing in this way is to ensure that the top plate crack between two adjacent boreholes can be completely connected, resulting in more thorough crack initiation in the top plate and achieving a better pressure relief effect.

[0147] In some embodiments, the method for fracturing the high-level top plate and / or low-level top plate in step S33 includes at least one of the following: placing a fracturing agent in the fracturing section of the borehole and sealing the borehole to induce a reaction of the fracturing agent to fracture the high-level top plate and / or low-level top plate; sealing the fracturing section of the borehole and continuously injecting high-pressure fluid into the fracturing section to fracture the high-level top plate and / or low-level top plate.

[0148] Specifically, the fracturing agent can be an explosive or a chemical agent that can expand in volume through a chemical reaction. The high-pressure fluid can be high-pressure water, etc.

[0149] In some embodiments, step S33 involves injecting high-pressure fluid into the fracturing zone within the borehole and fracturing the top plate using the high-pressure fluid, and step S33 includes the following steps:

[0150] S331: Divide the fracture-causing section of the borehole into multiple borehole segments along the borehole extension direction;

[0151] S332: Seal both ends of each borehole section in sequence and continuously inject high-pressure fluid so that the high-level top plate and / or low-level top plate can be layered along the borehole section and collapse in sequence.

[0152] Specifically, the fracture-causing zone of each borehole can be divided into 5 borehole segments. One borehole segment is first sealed, and high-pressure fluid is continuously introduced into that segment to allow the top plate fracture at the location of that borehole segment to be connected between two adjacent boreholes. The remaining borehole segments are treated in the same way.

[0153] In existing methods, the drilling-induced cracking zones in the roof are relatively concentrated, resulting in poor collapse performance due to the large thickness and good integrity of the roof. In this method, the borehole is divided into multiple drilling sections, and each section is sealed and fractured separately. The drilling-induced cracking zones are more dispersed, allowing the roof to collapse in layers with better collapse performance.

[0154] In some embodiments, step S2, determining the positions of the high-level roof and the low-level roof, includes the following steps: S21: Drilling boreholes above the coal seam of the working face for observation and taking core samples to determine the physical and mechanical parameters of each roof layer; S22: Performing theoretical calculations based on the physical and mechanical parameters of each roof layer to obtain position data one for the high-level roof and the low-level roof; S23: Collecting spatial location data of microseismic events occurring during the working face mining period to analyze and obtain position data two for the high-level roof and the low-level roof; S24: Comparing position data one and position data two, and determining the positions of the high-level roof and the low-level roof when the deviation between position data one and position data two meets a set standard. By combining position data one and position data two, the positions of the high-level roof and the low-level roof can be determined more accurately.

[0155] In other embodiments, the positions of the high-level top plate and the low-level top plate can be determined using only position data one or only position data two.

[0156] The following describes the method for stopping mining under equal pressure according to an embodiment of the present invention.

[0157] The method for stopping mining and equalizing pressure according to embodiments of the present invention includes the following steps:

[0158] Z1: The roof of the coal mine working face in the final mining stage is subjected to fracturing treatment according to the roof fracturing method in any of the above embodiments. Generally, the roof fracturing treatment in the final mining stage can be carried out when the coal body is mined to a distance of about 300 meters from the retreat roadway.

[0159] Z2: When the working face is in its final mining phase, mining is suspended after passing through fault line L2, and the roof behind fault line L2 is allowed to collapse. Since the high and low roof behind fault line L2 has already been fractured, and the collapse of the roof has a certain lag, the purpose of suspending mining is to wait for the roof to collapse fully, thereby reducing the occurrence of dynamic disasters.

[0160] Z3: Continue mining the remaining coal seam and pass through fault line L1.

[0161] Z4: Downhole equipment withdrawn from the working face.

[0162] After the coal body continues to be mined past fracture line L1, the mining speed can be accelerated to complete the extraction of the last coal body. Since the roof between fracture lines L1 and L2 has been fractured and can only serve as a temporary support, the remaining coal body must be mined quickly and the underground equipment must be removed in a timely manner to prevent the roof from collapsing and pressing on the underground equipment, making it impossible to remove the equipment.

[0163] The following describes the method for inducing roof cracking in the final stage of coal mine working face according to the present invention through Example 1.

[0164] This example describes the 30202 longwall mining face in a rockburst-prone mine. The 30202 longwall mining face is buried at a depth of approximately 647m. It is adjacent to the mined-out 30201 working face to the east and the designed 30203 working face with solid coal to the west. It is the first continuous working face in this mine.

[0165] Specific parameters of working face 30202: W 1. Take 5.4m W 2. Take 0.5m. W 3. Take 6.1m, D Take 24m, f 1. Take 25MPa f 2. Take 33MPa, γ Take 2.35×10⁴ N / m 3 , H Take 647m, h Take 5.68m, K Take 1.05, λ Take 0.5, C Take 5.5MPa, φ Take 32° σ Using a pressure of 25.41 MPa, the distance from the fracture line L1 to the non-production side of the retreat roadway was calculated using the above parameters. W The distance is 12m; the distance of the fracture line L2 from the non-production side of the retreat roadway. It cannot exceed 29.4m.

[0166] In this embodiment, hydraulic fracturing technology is used to pre-fracture the thick, hard overlying roof to relieve pressure and mitigate risks in the local mining area. Taking into account the occurrence of key strata, borehole observation results near the stop line, and high-energy microseismic events during the mining of the 30202 working face, the 3rd and 8th roof layers above the coal seam were identified as the target strata for hydraulic fracturing. The 3rd and 8th roof layers are fine-grained sandstone, corresponding to the low-lying and high-lying roof layers that play a major role in controlling roof collapse, respectively.

[0167] like Figures 7-10As shown, drilling operations were carried out on the 30202 working face. Based on the required fracturing layer and the actual downhole operating conditions, the following borehole layout was determined:

[0168] like Figure 8 As shown, borehole a has a depth of 35m and an inclination angle of 45°; borehole b has a depth of 50m and an inclination angle of 60°, with a distance of 10m between boreholes a and b. Boreholes a and b were constructed within the retreat roadway 3, and were respectively a shallow low-level borehole and a deep high-level borehole. The horizontal projections of the fracturing zones of boreholes a and b are both located behind the fracture line L1, and were used to fracturing the 3rd and 8th roof layers, respectively.

[0169] like Figure 9 As shown, borehole c has a depth of 30m and an inclination angle of 60°, with a spacing of 8m between boreholes. Borehole c is constructed within the return air roadway 4 and the adhesive transport roadway 5. The horizontal projection of the cracking section of borehole c is located between fracture lines L1 and L2, and it is used to crack the third-layer top slab.

[0170] like Figure 10 As shown, borehole d has a depth of 50m and an inclination angle of 60°. The spacing between boreholes d is 8m, and the spacing between boreholes c and d is also 8m. Borehole d is constructed within the return air roadway 4 and the adhesive transport roadway 5. The horizontal projection of the cracking section of borehole d is located behind the fracture line L2 and is used for cracking treatment of the 8th floor top slab.

[0171] According to downhole test results, a water injection pressure of 35 MPa and a fracturing time of 30 minutes can achieve fracture penetration. To promote the stratified fracturing of the thick, hard roof, the borehole was divided into five sections, each section was sealed, and five fracturing operations were performed on each borehole. Thus, the roof fracturing treatment for the final stage of the 30202 working face was completed.

[0172] The following uses stress monitoring data, microseismic monitoring data, and mine pressure monitoring data from the final mining stage of the working face to verify the hydraulic fracturing effect of the roof in the 30202 working face.

[0173] Stress monitoring data. Stress data from measuring points 4107# and 109# in the return air roadway of the 30202 working face were analyzed. Measuring point 107# is located in the non-fracturing section, and measuring point 109# is located in the fracturing section. Figure 11 , Figure 12 As shown, the stress in the coal seam of the hydraulically fractured section is significantly lower than that of the non-fracturing section, indicating that the roof collapse resistance of the hydraulically fractured section is good and the coal seam pressure relief effect is obvious.

[0174] Microseismic monitoring data and mine pressure monitoring data. For example... Figures 13-16 As shown, microseismic monitoring data from the 30202 working face were selected for analysis. Microseismic events after January 1st were statistically analyzed, with the mining range approximately 167m to 0m from the stop line. This range was further subdivided into three sections:

[0175] (1) From January 1 to 10, the mining range was approximately 167m to 86m away from the stop line. At this time, the retreat roadway had already been affected by the advance mining, but the peak of the advance support pressure had not yet reached the retreat roadway.

[0176] (2) From January 11 to 20, the mining range was approximately 86m to 32m from the stop line. According to the online monitoring of stress in the retreat roadway, the peak pressure of the working face advance support reached the retreat roadway at approximately 86m from the stop line, but this section had not yet entered the fracturing range.

[0177] (3) From January 21 to January 30, the mining range was approximately 32m to 0m away from the stop mining line. This section had entered the fracturing range.

[0178] A total of 1,681 microseismic events occurred in the first section, including 10 cubic events. During this phase, the microseismic events were mainly concentrated at the entrance of the No. 3 return air tunnel and in the area of ​​contour changes in the floor slab in front of the haulage roadway, with only a few events occurring at the entrance of the No. 2 return air tunnel and within the main retreat tunnel. This indicates that the thick, hard roof structure was relatively stable at this time, and the thick, hard roof was within an energy accumulation cycle. Comparing the on-site pressure step analysis, there were 17 continuous pressure strokes from January 5th to 7th and 12 continuous pressure strokes from January 9th to 10th, further confirming the integrity of the thick, hard roof structure.

[0179] A total of 1,587 microseismic events occurred in the second section, including 11 cubic events. These microseismic events began to cluster at the intersection of the retreat roadway and the return airway, exhibiting a high density and a clustered distribution. The affected area covered directly above the main and auxiliary retreat roadways, primarily releasing small amounts of energy. The number of microseismic events increased, and their concentration further increased, especially at distances of 86m to 35m from the stop line. However, overall, the energy of the microseismic events was not significantly higher than in the previous section, indicating that the energy accumulated in the thick, hard roof had not been effectively released. This suggests that although the width of the remaining coal pillar further narrowed, leading to a further concentration of mine pressure, the thick, hard roof structure was not fundamentally damaged, the energy storage structure remained, and the roof was still within its energy storage cycle. On-site pressure data showed 21 continuous pressure surges from January 11th to 15th. Compared to the previous section, the continuous pressure step was further increased, indicating a very high risk of impact behind the large-scale roof pressure.

[0180] A total of 1977 microseismic events occurred in the third section, including 16 cubic events and 3 events of magnitude 4 or higher. In this section, the number of low-energy microseismic events within 100m on the return air side decreased, and the release of microseismic energy was more concentrated. This indicates that the hydraulic fracturing technology on the roof destroyed the thick, rigid roof structure, allowing the accumulated energy to be released. On-site pressure data showed four continuous pressure cuts on January 20th and five on January 22nd. Compared to the first two stages, the area of ​​roof pressure within the fracturing section was significantly reduced, further proving that the hydraulic fracturing technology successfully destroyed the energy storage structure of the thick, rigid roof.

[0181] Three events of power 4 or higher occurred on January 23, 24, and 27, approximately 22m, 17m, and 17m from the stop-mining line, respectively. These events passed within the fracturing range of the high-level fracturing borehole D in the roof (D1 and D3 are located 24m from the stop-mining line in the return airway 4 and the conveyor roadway 5; D2 and D4 are located 32m from the stop-mining line in the return airway 4 and the conveyor roadway 5). Furthermore, the high-energy microseismic events on January 23, 24, and 27 occurred within the fracturing layer, indicating that the high-level fracturing borehole successfully controlled the distance between the fracture location of the old roof and the stop-mining line during the stop-mining adjustment period, thus achieving the goal of relieving pressure on the working face. Since then, energy release events have primarily been of power 3 or lower, further verifying that the low-level borehole disrupted the energy storage structure of the low-level roof.

[0182] like Figure 16 As shown, the support resistance data of the 30202 working face were selected for analysis. The average pressure step distance from January 5th to 22nd was statistically analyzed. The average pressure step distance in the non-fracturing section was 12.6 cuts, the average continuous pressure step distance was 9.2 cuts, and the average pressure step distance after entering the fracturing section was 10.5 cuts, while the average continuous pressure step distance was 4.5 cuts. The pressure step distance of the roof after entering the fracturing section and the continuous pressure step distance were significantly shortened, indicating that the double-end sealing multi-stage hydraulic fracturing technology successfully disrupted the structural integrity of the thick and hard roof.

[0183] In the description of this invention, it should be understood that the terms "center," "longitudinal," "lateral," "length," "width," "thickness," "upper," "lower," "front," "rear," "left," "right," "vertical," "horizontal," "top," "bottom," "inner," "outer," "clockwise," "counterclockwise," "axial," "radial," and "circumferential" indicate the orientation or positional relationship based on the orientation or positional relationship shown in the accompanying drawings. They are used only for the convenience of describing this invention and simplifying the description, and are not intended to indicate or imply that the device or element referred to must have a specific orientation, or be constructed and operated in a specific orientation. Therefore, they should not be construed as limitations on this invention.

[0184] Furthermore, the terms "first" and "second" are used for descriptive purposes only and should not be construed as indicating or implying relative importance or implicitly specifying the number of indicated technical features. Thus, a feature defined as "first" or "second" may explicitly or implicitly include at least one of that feature. In the description of this invention, "a plurality of" means at least two, such as two, three, etc., unless otherwise explicitly specified.

[0185] In the description of this specification, the references to terms such as "one embodiment," "some embodiments," "example," "specific example," or "some examples," etc., refer to specific features, structures, materials, or characteristics described in connection with that embodiment or example, which are included in at least one embodiment or example of the present invention. In this specification, the illustrative expressions of the above terms do not necessarily refer to the same embodiment or example. Furthermore, the specific features, structures, materials, or characteristics described may be combined in any suitable manner in one or more embodiments or examples. Moreover, without contradiction, those skilled in the art can combine and integrate the different embodiments or examples described in this specification, as well as the features of different embodiments or examples.

[0186] Although embodiments of the present invention have been shown and described above, it is understood that the above embodiments are exemplary and should not be construed as limiting the present invention. Those skilled in the art can make changes, modifications, substitutions and variations to the above embodiments within the scope of the present invention.

Claims

1. A method for inducing roof fracturing in a coal mine working face during the final mining stage, characterized in that, Includes the following steps: S1: Calculate the positions of fracture lines L1 and L2, where the distance between fracture line L1 and the retreat roadway is... W The distance between the fracture line L2 and the retreat roadway , in, W 1 represents the width of the retreat tunnel, in meters (m). W 2 represents the distance between the support frame and the stack frame, in meters; W 3 represents the control distance of the working face support, in meters; D The width of the remaining coal pillar, in meters; S2: Determine the positions of the high-level roof plate and the low-level roof plate that play a major role in controlling the roof collapse, wherein the height of the high-level roof plate from the working surface is greater than the height of the low-level roof plate from the working surface; S3: Perform cracking treatment on the low-position top plate and part of the high-position top plate located between the fracture lines L1 and L2, and perform cracking treatment on the low-position top plate and the high-position top plate located behind the fracture line L2.

2. The method for fracturing the roof of a coal mine working face in the final mining stage according to claim 1, characterized in that, In step S1, the width of the remaining coal pillar D Calculated using the following formula: = in, The average stress of the remaining coal pillar is given in kPa. The total load of the remaining coal pillar is kN; The critical width, in meters, required to maintain the stability of the remaining coal pillar. The sloping length is taken as 1m; The strength of the coal pillar is given in kPa. The uniaxial compressive strength of the cubic coal sample is given in kPa. The depth is measured in meters (m).

3. The method for fracturing the roof of a coal mine working face in the final mining stage according to claim 1, characterized in that, Step S3 includes the following steps: S31: Divide the roof above the coal seam in the retreat roadway into region A. The roof above the coal seam in the return air roadway, located between the fracture lines L1 and L2, is designated as region B1. The roof above the coal seam in the conveyor roadway, located between the fracture lines L1 and L2, is designated as region B2. The roof above the coal seam of the return air roadway, located behind the fracture line L2, is divided into region C1. The roof above the coal seam of the conveyor roadway, located behind the fracture line L2, is divided into region C2; S32: Within region A, boreholes are drilled from the retreat roadway to the roof above the coal face, such that the fracturing section of a portion of the borehole reaches the high roof behind the fracture line L1, and the fracturing section of a portion of the borehole reaches the low roof behind the fracture line L1. And / or, within regions B1 and B2, boreholes are drilled from the return air roadway and the conveyor roadway toward the roof above the coal seam of the working face, such that the fracturing zone of the boreholes reaches the lower roof. And / or, within the regions C1 and C2, boreholes are drilled from the return air roadway and the conveyor roadway toward the roof above the coal body of the working face, so that the fracturing section of the borehole reaches the low roof and the high roof. S33: Crack the high-level top plate and / or the low-level top plate by drilling holes in the regions A, B1, B2, C1, and C2.

4. The method for inducing roof fracturing in the final stage of coal mine working face according to claim 3, characterized in that, In step S32, the boreholes in region A include high-level deep holes and low-level shallow holes, and the high-level deep holes and low-level shallow holes are arranged alternately along the extension direction of the retreat roadway. The fracturing section of the high-level deep holes reaches the high-level roof, and the fracturing section of the low-level shallow holes reaches the low-level roof.

5. The method for inducing roof fracturing in the final stage of coal mine working face according to claim 3, characterized in that, In step S32, the drill holes in region C1 are arranged at intervals along the return air roadway, and the drill holes in region C2 are arranged at intervals along the extension direction of the adhesive conveying roadway, and the number of drill holes in regions C1 and C2 is 2-6.

6. The method for inducing roof fracturing in the final stage of coal mine working face according to claim 3, characterized in that, In step S22, the drilling spacing d Calculated using the following formula: in, r Let be the crack radius, in meters.

7. The method for inducing roof fracturing in the final stage of coal mine working face according to claim 3, characterized in that, The method for cracking the high-level top plate and / or the low-level top plate in step S33 includes at least one of the following: A fracturing agent is placed in the fracturing section of the borehole and the borehole is sealed to trigger a reaction of the fracturing agent to fracture the high-level top plate and / or the low-level top plate. The fractured section of the borehole is sealed, and high-pressure fluid is continuously injected into the fractured section to fracture the high-level top plate and / or the low-level top plate.

8. The method for inducing roof fracturing in the final stage of coal mine working face according to claim 7, characterized in that, In step S33, high-pressure fluid is used to cause cracks in the high-level top plate and / or the low-level top plate, and step S33 includes the following steps: S331: Divide the fracture-causing section of the borehole into multiple borehole segments along the borehole extension direction; S332: Seal both ends of each drilled section in sequence and continuously inject high-pressure fluid so that the high-level top plate and / or the low-level top plate can be layered along the drilled section and collapse in sequence.

9. The method for inducing roof fracturing in the final stage of coal mine working face according to claim 1, characterized in that, In step S2, determining the positions of the high-level top plate and the low-level top plate includes the following steps: S21: Drill holes above the coal seam of the working face to observe and take rock cores to determine the physical and mechanical parameters of each layer of the roof; S22: Based on the physical and mechanical parameters of each top plate, perform theoretical calculations to obtain the position data of the high-level top plate and the low-level top plate; S23: Collect spatial location data of micro-seismic events that occur during the mining of the working face, and analyze the location data of the high-level roof and the low-level roof to obtain the second set of location data; S24: Compare the position data one and the position data two. When the deviation between the position data one and the position data two meets the set standard, determine the position of the high-position top plate and the low-position top plate.

10. A method for stopping mining under equal pressure based on the roof fracturing method of the final stage coal mine working face according to any one of claims 1-9, characterized in that, Includes the following steps: Z1: The roof of the coal mine working face in the final mining stage is subjected to fracturing treatment according to any one of claims 1-9. Z2: Conduct the final mining operation at the working face. After mining passes through the fracture line L2, stop mining and wait for the roof behind the fracture line L2 to collapse. Z3: Continue mining the remaining coal seam and pass through fracture line L1; Z4: Remove downhole equipment from the working face.

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