Methods for controlling the fracturing of the hard roof in horizontal segmented fully mechanized longwall mining of steeply inclined thick coal seams

Abstract

This invention discloses a method for controlling the fracturing of the hard roof in horizontal segmented fully mechanized longwall mining of steeply inclined thick coal seams. First, the extent of the roof collapse zone is determined. Then, roof boreholes are drilled along the edge and inside of the defined roof collapse zone in the transport and return airways, and in-seam boreholes are drilled into the coal seam. Finally, a multi-hole simultaneous fracturing method is used on the roof boreholes drilled along the edge of the roof collapse zone to directionally sever the connection between the roof collapse zone and the underlying rock strata. For the remaining boreholes, a pulse fracturing method is used to induce fatigue impact damage in the coal and rock mass, forming a complex fracture network within it. This effectively weakens the hard roof and the underlying top coal. When the fracturing zone enters the goaf, the top coal and roof are fully fractured and collapsed under mine pressure, and the collapsed roof promptly fills the goaf. This method eliminates the safety hazards caused by the accumulation of large amounts of gas in cavities formed in the upper goaf, and also improves the recovery rate of top coal and reduces resource waste.

Description

Technical Field

[0001] This invention relates to the field of coal mining technology, specifically to a method for controlling the fracturing of the hard roof in horizontal segmented fully mechanized longwall mining of steeply inclined thick coal seams. Background Technology

[0002] Steeply dipping coal seams refer to coal seams with a dip angle greater than 45°, accounting for approximately 17% of total coal reserves. Especially in western regions, more than half of the mines are mining steeply dipping coal seams. When steeply dipping coal seams are thick, horizontal segmented fully mechanized longwall mining is generally employed. However, the difficulty in collapsing the hard roof during horizontal segmented fully mechanized longwall mining of thick, steeply dipping coal seams has always been a challenge, posing serious safety hazards to mine production.

[0003] The main reasons why the hard roof is difficult to collapse in the horizontal segmented fully mechanized longwall mining of steeply inclined thick coal seams are as follows: (1) Compared with near-horizontal and gently inclined coal seam mining, the migration, collapse and stress distribution characteristics of the overlying strata in the steeply inclined coal seam working face are significantly different. The tangential component of the load on the overlying strata is greater than the normal component along the strata, and the mine pressure is not obvious; (2) In the horizontal segmented fully mechanized longwall mining of steeply inclined thick coal seams, the working face is short and the mine pressure is not obvious; (3) When the roof is a rock stratum with characteristics such as large layer thickness, good integrity and high strength, it increases the difficulty of timely collapse of the roof in the goaf. In addition to the dynamic disaster caused by the sudden collapse of the large area of ​​the goaf roof, the formation of a large area of ​​the goaf roof will cause the accumulation of gas in the goaf; the mine pressure is not obvious and the top coal is not broken enough, resulting in a lot of residual coal in the goaf. If these problems are not addressed in a timely manner, improper provision of coal pillars in the lower section during mining can easily lead to communication between the upper and lower goaf areas, air leakage in the upper goaf area, spontaneous combustion of coal left in the upper goaf area, igniting or detonating a large amount of gas accumulated in the upper goaf area, causing a major safety accident.

[0004] Traditional methods for controlling the hard roof of steeply dipping coal seams involve explosive blasting. However, explosive blasting weakens the rock strata, complicating safety management; the large amounts of harmful gases such as CO generated instantaneously during large-scale blasting significantly impact mine ventilation safety management; for high-gas mines, explosive blasting is unsuitable due to the risk of gas explosions induced by blasting sparks; the small effective range of a single blast hole necessitates large quantities of explosives and detonators, resulting in high economic costs; and during deep-hole blasting, the confining pressure limits the range of crack formation and the rock-breaking effect. Therefore, there is an urgent need to propose a safe and efficient method for controlling the hard roof in horizontal segmented fully mechanized longwall mining of steeply dipping thick coal seams. Summary of the Invention

[0005] The purpose of this invention is to provide a method for controlling the fracturing of the hard roof in horizontal segmented fully mechanized longwall mining of steeply inclined thick coal seams. This method can effectively ensure the timely collapse of the hard roof and the filling of the goaf during the horizontal segmented fully mechanized longwall mining of steeply inclined thick coal seams, thereby ensuring safe and efficient mining.

[0006] To achieve the above objectives, the present invention provides a method for controlling fracturing of the hard roof in horizontal segmented fully mechanized longwall mining of steeply inclined thick coal seams, comprising the following steps:

[0007] S1. Determine the extent of the roof collapse zone;

[0008] S2. On one sidewall of the transport roadway, perpendicular to the direction of the transport roadway and along the connection between the roof collapse area and the lower rock strata, a first row of long roof boreholes is drilled; on the roof of the transport roadway, perpendicular to the direction of the transport roadway and along the diagonal of the roof collapse area, a second row of long roof boreholes is drilled; on the roof of the transport roadway, perpendicular to the direction of the transport roadway and along the angle bisector of the angle formed by the first and second rows of long roof boreholes, a third row of long roof boreholes is drilled; on the other sidewall of the transport roadway, a row of coal seam in-seam boreholes is drilled, the length of which is determined according to the coal discharge height; all four rows of boreholes are linearly arranged along the direction of the transport roadway.

[0009] S3. On the top plate of the return air roadway, perpendicular to the direction of the return air roadway and parallel to the direction of the long boreholes of the second row of top plates of the transport roadway, a first row of short boreholes is drilled, with the final hole located at the connection surface between the collapsed area of ​​the top plate and the upper rock strata; on the top plate of the return air roadway, perpendicular to the direction of the return air roadway and towards the projection point of the midpoint of the long boreholes of the second row of top plates of the transport roadway at this cross section, a second row of short boreholes is drilled, with the final hole located at the projection point of the midpoint of the long boreholes of the second row of top plates of the transport roadway at this cross section; both the first row of short boreholes and the second row of short boreholes are linearly arranged along the direction of the return air roadway.

[0010] S4. The first row of long boreholes in the transport roadway is subjected to multi-hole simultaneous fracturing, so that the cracks extend along the direction of the borehole connection line and directionally cut off the connection between the collapse area of ​​the roof and the lower rock layer.

[0011] S5 and the remaining boreholes all use pulse fracturing to induce fatigue impact damage in the coal and rock mass, forming a complex network of fractures within it, which effectively weakens the hard roof of the coal and rock mass and the top coal below it.

[0012] Furthermore, the cross-section of the roof collapse zone is similar to that of a parallelogram, and the width of the roof collapse zone is determined by the following formula:

[0013]

[0014] Where W is the width of the roof collapse zone; H0 is the segment height of the coal mining face; α is the coal seam dip angle; the minimum collapse height of the roof collapse zone (10) is determined by the following formula:

[0015]

[0016] Where H is the minimum collapse height that can fill the goaf; h is the coal seam thickness; and k is the fracture expansion coefficient of the collapsed rock strata.

[0017] The length of the roof collapse zone is determined by the following formula:

[0018] L = L0

[0019] Where L is the length of the roof collapse zone; L0 is the length of the mineable strike of the working face.

[0020] Furthermore, the first row of long boreholes in the transport roadway is located at the right edge of the roof collapse zone; the final positions of the second row of long boreholes and the third row of long boreholes in the transport roadway are located at the upper edge of the roof collapse zone; the final positions of the coal seam in-seam boreholes in the transport roadway and the first row of short boreholes in the return air roadway are located at the left edge of the roof collapse zone; the interface between the coal seam and the roof is the lower edge of the roof collapse zone.

[0021] Furthermore, if drilling and fracturing the second row of long roof holes, the third row of long roof holes, the coal seam in-seam holes, and the first row of short roof holes and the second row of short roof holes in the return air roadway can cause the roof to collapse in a timely manner, then the first row of long roof holes in the transport roadway is not required.

[0022] The beneficial effects of this invention are as follows: This method enables the roof of the goaf to collapse and fill the goaf in a timely manner after horizontal segmented fully mechanized mining of steeply inclined thick coal seams, avoiding dynamic disasters caused by sudden collapse of large-area suspended roofs, and eliminating the safety hazards caused by the formation of cavities and the accumulation of large amounts of gas in the upper goaf. Furthermore, it can improve the recovery rate of top coal and reduce resource waste. Since drilling and fracturing are carried out in the transport roadway and return airway, it does not affect the normal mining of the working face. Compared with traditional explosive blasting methods, the main advantages of this method are that the hydraulic fracturing control method causes less disturbance to the surrounding rock of the mining area, produces no harmful gases, and is simple to construct with a smaller workload. Attached Figure Description

[0023] Figure 1 Schematic diagram of tunnel layout;

[0024] Figure 2 Schematic diagram of the overhanging roof formed in the goaf after mining;

[0025] Figure 3 Schematic diagram of the cavity formed in the upper section of the goaf;

[0026] Figure 4 Diagram illustrating the causes of the disaster;

[0027] Figure 5 A schematic diagram showing the delineation of the roof collapse zone, borehole layout, fracturing, and the effects of roof collapse.

[0028] In the diagram, 1-floor, 2-coal seam, 3-roof, 4-return airway, 5-transport roadway, 6-working face, 7-goaf, 8-broken rock, 9-collapsed rock strata, 10-roof collapse area, 11-left edge of roof collapse area, 12-upper edge of roof collapse area, 13-right edge of roof collapse area, 14-lower edge of roof collapse area, 15-curved rock strata, 16-fresh airflow, 17-coal pillar, 18-air leakage, 19-... - Residual coal, 20- Spontaneous combustion of residual coal, 21- Gas, 22- Gas explosion, 23- Upper section goaf, 24- Long borehole in the first row of roof, 25- Long borehole in the second row of roof, 26- Long borehole in the third row of roof, 27- Coal seam in-seam borehole, 28- Short borehole in the first row of roof, 29- Short borehole in the second row of roof, 30- Pulse-pressure fracture network, 31- Minimum collapse height that can fill the goaf, 32- Low-level roof, 33- High-level roof. Detailed Implementation

[0029] The present invention will be further described in detail below with reference to the accompanying drawings and specific embodiments.

[0030] like Figure 1 As shown, the coal seam mined in the No. 4 coal seam of the East 2 working face of a high-gas mine is 6m thick with a dip angle of 45°, and is classified as a spontaneously combustible coal seam. The immediate roof is a coarse sandstone with an average thickness of 16m, exhibiting good integrity without obvious joints or fissures, and a uniaxial compressive strength of 99MPa. The mineable strike length of the working face is approximately 200m, with an inclination length of 14m. The working face adopts a horizontal segmented fully mechanized top-coal caving mining method, with a segment height of approximately 17m, including a recovery height of 2.5m. The return air roadway 4 and haulage roadway 5 are located on either side of the working face 6, both arranged along the strike direction of coal seam 2. Below coal seam 2 is the floor 1, and above it is the roof 3.

[0031] like Figure 2 As shown, during the horizontal segmented fully mechanized longwall mining of steeply inclined thick coal seams, when the roof of the coal seam is relatively hard, the roof 3 of the working face 6, which is closer to the goaf 7, will not collapse.

[0032] like Figure 3As shown, as the working face advances, the roof of the goaf, which is relatively far from the working face, begins to collapse. Due to the large angle of inclination, the falling gangue slides down along the inclined direction to fill the lower goaf, while the upper goaf forms a cavity.

[0033] like Figure 3 As shown, the sliding backfilling of collapsed rock prevents the direct roof of the lower inclined section of the stope from moving downwards sufficiently. Usually, only the surface rock strata collapse, forming broken rock blocks 8. The layers above these blocks are all fractured to the size allowed by their respective lithologies and arranged neatly, forming collapsed rock strata 9. These broken rock blocks 8 provide some support to the overlying rock strata, causing the overlying rock strata to only bend and sink without fracturing, forming bent rock strata 15, thus creating a stable structure.

[0034] like Figure 4 As shown, in addition to the dynamic disaster caused by the sudden collapse of a large area of ​​overhanging roof in the goaf, the formation of a large area of ​​overhanging roof in the goaf can cause gas accumulation in the goaf; the mine pressure is not obvious and the top coal is not sufficiently broken, resulting in a large amount of residual coal 19 in the goaf. If these problems are not dealt with in time, when the lower section is mined, if the section coal pillar 17 is not properly reserved, it is easy to cause the goaf of the upper and lower sections to be connected. Fresh air 16 leaks into the upper section goaf through the coal pillar 17, causing the residual coal 23 in the upper section to spontaneously combust 20, igniting or detonating a large amount of gas 21 accumulated in the upper section goaf, forming a gas explosion 22, which can easily cause a major safety accident.

[0035] To address the aforementioned issues, a combined approach of directional fracturing to cut through the lower strata of the stope and pulse fracturing to weaken the hard top coal roof is proposed. The specific steps are as follows:

[0036] like Figure 5 As shown in section AA, the extent of the roof collapse zone 10 is determined. The cross-section of the roof collapse zone is similar to that of a parallelogram, and the width of the roof collapse zone is determined by the following formula:

[0037]

[0038] Where W is the width of the roof collapse zone; H0 is the segment height of the coal face; and α is the dip angle of the coal seam. Therefore, the width of the roof collapse zone is determined to be 24.7m.

[0039] The minimum collapse height of the roof collapse zone 10 is determined by the following formula:

[0040]

[0041] Where H is the minimum collapse height of 31 m that can fill the goaf; h is the coal seam thickness; and k is the fracture expansion coefficient of the collapsed rock strata. Therefore, the height of the roof collapse zone is determined to be 20 m.

[0042] The length of the roof collapse zone is determined by the following formula:

[0043] L = L0

[0044] Where L is the length of the roof collapse zone; L0 is the length of the mineable strike of the working face. Therefore, the length of the roof collapse zone is determined to be 200m. The sum of the bottom roof 31 and the upper roof 32 is the minimum collapse height 31 that can fill the goaf.

[0045] like Figure 5 As shown in the BB cross-section, on one sidewall of transport roadway 5, perpendicular to the direction of transport roadway 5 and along the connection between the roof collapse zone 10 and the lower rock strata, a first row of long boreholes 24 are drilled. The boreholes are 21.4m long, with an inclination angle of 62° and a spacing of 10m. The roof collapse zone is at a certain angle to the horizontal direction, i.e., it is in an inclined state. The lower rock strata are those connected to the bottom edge of the roof collapse zone 10, and are also in an inclined state. After moving 2m outward, a second row of long boreholes 25 are drilled on the roof of transport roadway 5, perpendicular to the direction of transport roadway 5 and along the diagonal of the roof collapse zone 10. The boreholes are 35m long, with an inclination angle of 84° and a spacing of 10m. After moving another 2m outward, a third row of long boreholes 26 are drilled on the roof of transport roadway 5, perpendicular to the direction of transport roadway 5 and along the angle bisector of the angle formed by the first row of long boreholes 24 and the second row of long boreholes 25. The boreholes are 24m long, with an inclination angle of 76° and a spacing of 10m. After moving another 2m outward, a row of coal seam in-seam boreholes 27 are drilled on the other sidewall of transport roadway 5. The length of the coal seam in-seam boreholes 27 is determined according to the coal discharge height, with a length of 21.5m, an inclination angle of 45°, and a spacing of 10m.

[0046] like Figure 5 As shown in the BB view, on the top plate of the return air roadway 4, perpendicular to the direction of the return air roadway 4 and parallel to the direction of the second row of long boreholes 25 in the transport roadway 5, a first row of short boreholes 28 are drilled. The final borehole is located at the junction of the collapse zone 10 and the upper rock strata (located above the inclined surface of the collapse zone 10, connecting to the left edge of the collapse zone 10). The borehole length is 20.2m, the borehole inclination angle is 86°, and the borehole spacing is 10m. Moving outwards by 2m, on the top plate of the return air roadway 4, perpendicular to the direction of the return air roadway 4, towards the projection point of the midpoint of the second row of long boreholes 25 in the transport roadway 5 at this section, a second row of short boreholes 29 are drilled. The final borehole is located at the projection point of the midpoint of the second row of long boreholes 25 in the transport roadway 5 at this section. The borehole length is 15m, the borehole inclination angle is 65°, and the borehole spacing is 10m.

[0047] like Figure 5As shown, the first row of long roof boreholes 24 in the transport roadway 5 is located at the right edge 13 of the roof collapse zone; the final positions of the second row of long roof boreholes 25 and the third row of long roof boreholes 26 in the transport roadway 5 are located at the upper edge 12 of the roof collapse zone; the final positions of the coal seam in-seam boreholes in the transport roadway and the first row of short roof boreholes in the return air roadway are located at the left edge 11 of the roof collapse zone. The interface between the coal seam and the roof is the lower edge 14 of the roof collapse zone.

[0048] like Figure 5 As shown in the CC view, the first row of long boreholes 24 in the transport roadway 5 employs a multi-hole simultaneous fracturing method, causing the fractures to extend along the borehole connection line and directionally severing the connection between the roof collapse zone and the underlying rock strata. The remaining boreholes employ pulse fracturing to weaken the hard roof and the underlying coal seam within the roof collapse zone. The pumping flow rate of the pulse fracturing is a high-frequency, periodic fluctuation in the form of pulse waves, causing fatigue impact damage to the coal and rock mass and forming a complex network of pulse fractures 30 within it, thereby fully weakening the coal and rock mass.

[0049] like Figure 5 As shown in the DD view, when the fracturing zone enters the goaf 7, the top coal roof can be fully broken and collapsed under the action of mine pressure, and the collapsed roof can fill the goaf 7 in time.

[0050] In addition, if drilling and fracturing the second row of long roof holes 25, the third row of long roof holes 26, the coal seam in-seam holes 27, and the first row of short roof holes 28 and the second row of short roof holes 29 of the return air roadway 4 can cause the roof 3 to collapse in time, then it is not necessary to drill the first row of long roof holes 24 of the transport roadway 4.

[0051] The embodiments of the present invention have been described in detail above with reference to the accompanying drawings. However, the present invention is not limited thereto. Various changes that can be made within the scope of knowledge possessed by those skilled in the art without departing from the spirit of the present invention are all within the protection scope of the claims of the present invention.

Claims

1. A method for controlling fracturing of the hard roof in horizontal segmented fully mechanized longwall mining of steeply inclined thick coal seams, characterized in that, Includes the following steps: S1. Determine the extent of the roof collapse zone (10), the cross-section of which is similar to a parallelogram, and the width of which is determined by the following formula: ， in, W The width of the roof collapse zone; H 0 represents the segment height of the coal mining face; a The dip angle of the coal seam; The minimum collapse height of the roof collapse zone (10) is determined by the following formula: ， in, H The minimum collapse height required to fill the goaf; h Coal seam thickness; k The coefficient of fracturing of the collapsed rock strata; The length of the roof collapse zone is determined by the following formula: L = L 0 in, L The length of the roof collapse zone; L 0 represents the exploitable length of the working face; S2. On one side wall of the transport roadway (5), perpendicular to the direction of the transport roadway (5) and along the connection between the roof collapse area (10) and the lower rock strata, a first row of roof long boreholes (24) is drilled; on the roof of the transport roadway (5), perpendicular to the direction of the transport roadway (5) and along the diagonal of the roof collapse area (10), a second row of roof long boreholes (25) is drilled; on the roof of the transport roadway (5), perpendicular to the direction of the transport roadway (5) and along the bisector of the angle formed by the first row of roof long boreholes (24) and the second row of roof long boreholes (25), a third row of roof long boreholes (26) is drilled; on the other side wall of the transport roadway (5), a row of coal seam in-seam boreholes (27) is drilled, the length of which is determined according to the coal discharge height; all four rows of boreholes are arranged linearly along the direction of the transport roadway (5). S3. On the top plate of the return air roadway (4), perpendicular to the direction of the return air roadway (4) and parallel to the direction of the second row of long boreholes (25) of the transport roadway (5), a first row of short boreholes (28) is drilled, with the final hole position located on the connection surface between the collapse zone (10) of the top plate and the upper rock layer; on the top plate of the return air roadway (4), perpendicular to the direction of the return air roadway (4), and towards the projection point of the midpoint of the second row of long boreholes (25) of the transport roadway (5) at this section, a second row of short boreholes (29) is drilled, with the final hole position located at the projection point of the midpoint of the second row of long boreholes (25) of the transport roadway (5) at this section; both the first row of short boreholes (28) and the second row of short boreholes (29) are linearly arranged along the direction of the return air roadway (4); The first row of long boreholes (24) in the transport roadway (5) is located at the right edge of the roof collapse zone (10); the final positions of the second row of long boreholes (25) and the third row of long boreholes (26) in the transport roadway (5) are located at the upper edge of the roof collapse zone (10); the final positions of the coal seam in-seam boreholes (27) in the transport roadway (5) and the first row of short boreholes (28) in the return air roadway (4) are located at the left edge of the roof collapse zone (10); the interface between the coal seam and the roof is the lower edge of the roof collapse zone (10). S4. The first row of long boreholes (24) of the transport roadway (5) adopts a multi-hole simultaneous fracturing method to make the cracks extend along the direction of the borehole connection line and directionally cut off the connection between the top plate collapse area (10) and the lower rock layer. S5. The remaining boreholes all use pulse fracturing to cause fatigue impact damage to the coal and rock mass, forming a complex network of fractures inside, which fully weakens the hard roof of the coal and rock mass and the top coal below it.

2. The method for controlling fracturing of the hard roof in horizontal segmented fully mechanized longwall mining of steeply inclined thick coal seams according to claim 1, characterized in that, When drilling and fracturing the second row of roof long boreholes (25), the third row of roof long boreholes (26), the coal seam in-seam boreholes (27), and the first row of roof short boreholes (28) and the second row of roof short boreholes (29) of the return air roadway (4) can cause the roof to collapse in time, the first row of roof long boreholes (24) of the transport roadway (5) is not drilled.

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