Far-near field collaborative tip-cutting method

Abstract

The application discloses a far-near field cooperative roof cutting method and relates to the technical field of roof cutting. It comprises the following steps: using single hydraulic prop to cooperate with metal hinged roof beam and cross beam to pre-support the roof of the roadway, wherein the pre-roadway reinforcement is performed before the working face is mined, and the length of the pre-roadway reinforcement of the roadway to be left is appropriately lengthened, considering that the directional pre-splitting cutting hole is better in cutting seam outside the pre-pressure range. The application has the beneficial effects that: when in use, the pre-splitting roof cutting and the forced roof falling of the goaf block most of the mining pressure transmission to the roof of the roadway to be left, and the falling gangue supports the roof of the roadway to be left, thereby ensuring the stability of the roof of the roadway to be left along the goaf, and auxiliary support is performed during the roof cutting, so as to prevent the deformation of the roadway to be left from being too large and affecting the roadway to be left.

Description

Technical Field

[0001] This invention relates to the field of top cutting technology, and in particular to a far-field and near-field coordinated top cutting method. Background Technology

[0002] Currently, underground goaf retention mainly involves drilling holes along the design line in the roof of the retention roadway, using directional shaped charge blasting or hydraulic fracturing to pre-fracture along the design line in the direction of the roadway, and then using artificial materials to fill and support the roadway sides or setting up a rock retaining device to allow the collapsed rock to gradually accumulate and connect with the roof to relieve the pressure on the roof of the retention roadway. However, when the roof is a thick and hard rock layer, even if the roof can be cut open by simply cutting the roof, the roof is not easy to collapse in the goaf, and it is impossible to use the collapsed rock to build up support for the roof. When using artificial materials for filling, the filling material is not strong enough to support the pressure of the thick and hard roof, causing the filling wall to fail and the retention roadway to fail. Moreover, when the roof is a thick and hard rock layer, when cutting the roof by cutting the roof, the pre-fracture by blasting or the fracture of the roof will apply a large pressure to the upper part of the support structure, which will cause deformation of the retention roadway. Severe deformation of the retention roadway will cause significant losses and impacts. Summary of the Invention

[0003] The technical problem to be solved by this invention is to provide a far-field and near-field coordinated roof cutting method to address the aforementioned technical deficiencies. In use, pre-splitting roof cutting and forced roof caving in the goaf block most of the mining pressure from the mining area from being transmitted to the roof of the roadway, and the falling gangue supports the roof of the roadway, ensuring the stability of the roadway roof along the goaf. Furthermore, auxiliary support is provided during roof cutting to prevent excessive deformation during the roadway retention process from causing adverse effects.

[0004] The technical solution adopted in this invention is: to provide a far- and near-field coordinated top-cutting method, comprising the following steps:

[0005] S1: Use single hydraulic props in conjunction with metal articulated top beams and cross beams to reinforce and support the roof of the roadway in advance. The reinforcement of the roadway in advance should be carried out before the working face is mined. Considering that it is better to make cracks outside the advance pressure range of the directional pre-splitting blast holes, the length of the reinforcement of the roadway in advance should be appropriately increased to 20-25m. That is, the length of the reinforcement of the roadway outside the coal wall of the working face is 20-25m. As the working face is mined, the construction is continuously extended outward, keeping the minimum length not less than 20m.

[0006] S2: Construct roof-cutting blast holes in the roadway roof. The blast holes are arranged in a row along the roadway direction, with the openings of all holes on a straight line. The holes are driven towards the coal face of the working face. The drilling angle of the blast holes is 10-15° with the vertical line, the depth is 7-15m, the diameter is 50-70mm, the hole spacing is 0.5-1m, the charge per hole is 4-10kg, and the construction position is 2.5-5m away from the coal face of the roadway.

[0007] S3: In roadways where roadway retention is required, move the support mechanism to a range of 10-25m from the coal face. This range of 10-25m from the coal face is not only the range of advanced directional pre-splitting, but also the range of advanced support pressure of the working face. The roadway roof is supported and held in place by the support mechanism. The support position is the end of the roof that is close to the working face.

[0008] S4: In roadways where roadway retention is required, advance directional pre-splitting blasting is carried out within a range of 18-25m from the coal wall. Before construction, it is confirmed that the reinforcement quality of the advance roadway is intact. Mining-grade Class III emulsion explosive is placed in the shaped charge tube and connected to the detonation device. The shaped charge tube with charge is installed in the completed blasting borehole. The detonator leads are straightened. Then the borehole is sealed, the wire is connected, and the blasting is initiated to carry out directional pre-splitting, so that cracks appear in the roof along the edge of the designed roadway retention. Then the blasting effect is checked.

[0009] S5: When the roof is a thick, hard rock layer, even if simple roof cutting and jointing can cut through the roof, it is not easy for the roof to collapse in the goaf. In this case, forced roof caving is required. Bidirectional forced roof caving is adopted, that is, forced roof caving blast holes are drilled into the roof of the working face from both roadways. Each roadway is set up with 3 sets of holes in a cycle. The 3 sets of holes are numbered A, B and C respectively. They are all oriented towards the goaf and form angles of 70°, 45° and 10° with the roadway in the plane, that is, 20°, 45° and 80° with the vertical lines of the two roadways. Group A is drilled with 3 holes, numbered A1, A2 and A2. Group A3 has angles of 10°, 15°, and 20° with the roof, respectively. Group B has two holes, numbered B1 and B2, with angles of 23° and 35° with the roof, respectively. Group C holes are set differently depending on the destination of the goaf retaining roadway. For roadways without goaf retaining roadway, Group C holes are set as two holes, numbered C1 and C2, with angles of 20° and 30° with the roof, respectively. For roadways with goaf retaining roadway, Group C holes are set as three holes, numbered C1, C2, and C3, with angles of 20°, 30°, and 40° with the roof, respectively.

[0010] S6: Continuously charge explosives in holes of groups A, B, and C, leaving 2-3m uncharged at the hole opening. Then seal the holes, connect the wires, and detonate. Forced roof caving blasting causes the roof of the goaf to collapse, increasing the small-angle blasting along the roadway direction and the goaf, i.e., densifying the holes of group C. Utilize the pre-splitting blasting cut as a free surface to fully blast and break up the rock mass between the holes of group C and the pre-splitting cut, inducing it to collapse and accumulate to form a gangue belt, thus playing a supporting role during the collapse or subsidence of the overlying strata. Charge, seal, and connect the explosives just before the forced roof caving blast holes enter the goaf. Blasting is carried out as soon as they enter the goaf. Before blasting, strengthen the support within 20m of the blasting section to prevent blasting vibration from damaging some roadways that need maintenance. After blasting, check the blasting effect of forced roof caving to see if the goaf has collapsed and whether the roof of the remaining roadway is stable.

[0011] S7: The technical parameters for forced roof caving need to be adjusted according to the specific geological conditions of the coal mining face. The main parameters for adjustment are the hardness of the rock strata, the integrity of the rock strata, and the minimum resistance line radius of the blasting. When the hardness and integrity of the roof rock strata are high and the minimum resistance line radius of the blasting is small, the included angle between holes in the same group should be reduced and the number of holes should be increased appropriately. Conversely, the included angle between holes should be increased and the number of holes should be reduced appropriately.

[0012] The S3 support mechanism includes an L-shaped support plate; multiple connecting frames are evenly distributed at the lower end of the support plate; a support block is installed on the support plate between two adjacent connecting frames; a support device is provided at the lower end of each connecting frame; the support device includes an A electric push rod; the output end of the upper end of the A electric push rod is rotatably connected to a connector via a rotating shaft, and the connector and the connecting frame are engaged; an inclined B electric push rod is fixedly connected to the front side of the A electric push rod; pressure sensors are fixedly connected to the lower ends of both the A and B electric push rods, and the display screens of the pressure sensors are set on the corresponding A and B electric push rods; a fixed base is fixedly connected to the lower end of each pressure sensor; a fixed tooth is fixedly connected to the lower end of each fixed base; a drive bar is slidably connected to both ends of the A electric push rod; the upper end of each drive bar is fixedly connected to a connector; and a moving device is fixedly connected to the lower end of each drive bar.

[0013] To further optimize this technical solution, a connecting frame for a near-field and far-field coordinated top cutting method is an L-shaped connecting frame, which is fixedly connected to the support plate; the lower end of the connecting frame has a connecting port that engages with the connecting head.

[0014] To further optimize this technical solution, a method for coordinated near-field and far-field top cutting has a connecting groove at the lower end of each support block that matches a single hydraulic prop.

[0015] To further optimize this technical solution, a support block for a near-field and far-field coordinated roof cutting method is divided into support block A and support block B; the upper end face of support block A is located at the upper end of the inner side of the support plate; the upper end face and rear end face of support block B are fitted with the inner side of the support plate at an angle, and the connecting groove of support block B is set at an inclination.

[0016] To further optimize this technical solution, in a far-field and near-field coordinated roof cutting method, two mounting holes are opened on each of the two adjacent connecting frames on the support plate; the two mounting holes correspond to support block A and support block B, respectively.

[0017] To further optimize this technical solution, in a far- and near-field coordinated top cutting method, the support block has threaded holes corresponding to the mounting holes in the connecting groove; the mounting holes and threaded holes are connected by bolts; the lower end of the threaded hole is higher than the lower end of the bolt.

[0018] To further optimize this technical solution, a mobile device for a near-field and far-field coordinated roof cutting method includes a semi-circular tubular wheel seat; the upper end of the wheel seat is fixedly connected to the drive bar, and the wheel seat corresponds to and cooperates with the A support block; the lower end of the wheel seat is fixedly connected to a moving wheel.

[0019] To further optimize this technical solution, a self-locking omnidirectional wheel is used as the moving wheel in a far-field and near-field coordinated top cutting method.

[0020] Compared with traditional top-cutting methods, the advantages of this invention are as follows:

[0021] 1. By pre-splitting the roof and the improved forced roof caving in the goaf, the transmission of dynamic pressure from the goaf to the roadway area is blocked, and the roof of the roadway on the side of the goaf area collapses to support the overlying rock mass, relieving the pressure on the roof of the roadway and ensuring the stability of the roadway roof along the goaf.

[0022] 2. The support plate, A electric push rod, connector, connecting frame, connecting port, fixed base, fixed teeth and B electric push rod are used together to provide vertical and inclined support, resulting in better support effect;

[0023] 3. The combination of A electric push rod, B electric push rod, pressure sensor and display screen, after detecting the pressure, can be used in conjunction with advanced pre-splitting for safer auxiliary support to prevent excessive deformation of the roadway from causing an impact.

[0024] 4. It is convenient to adjust and reinforce the support according to different situations, making it easier to use; it is also easy to move and better used for advanced pre-splitting support. Attached Figure Description

[0025] Figure 1 This is an internal cross-sectional view of the advanced roadway reinforcement support of the present invention;

[0026] Figure 2 This is a plan view of the advanced roadway reinforcement support section of the present invention;

[0027] Figure 3 This is a plan view of the forced caving borehole of the present invention;

[0028] Figure 4 This is a cross-sectional view of three sets of holes for the forced top-caving blast holes of the present invention;

[0029] Figure 5 This is a plan view of the forced top-down structure before the present invention is completed;

[0030] Figure 6 This is a plan view of the forced top release according to the present invention;

[0031] Figure 7 This is a schematic diagram of the support mechanism of the present invention;

[0032] Figure 8This is a partial structural diagram of the support mechanism of the present invention;

[0033] Figure 9 This is a schematic diagram showing the structural changes of the support device of the present invention;

[0034] Figure 10 This is a schematic diagram of the support mechanism of the present invention from another perspective.

[0035] Figure 11 This is a partial exploded view of the support block of the present invention;

[0036] In the diagram, 1. Single hydraulic prop; 2. Metal articulated top beam; 3. Cross beam; 4. Top cutting blast hole; 5. Anchor cable; 6. Support plate; 7. Connecting frame; 8. Support block; 9. Electric push rod A; 10. Connector; 11. Electric push rod B; 12. Pressure sensor; 13. Display screen; 14. Fixed base; 15. Fixed tooth; 16. Drive bar; 17. Connection port; 18. Connection groove; 19. Support block A; 20. Support block B; 21. Mounting hole; 22. Threaded hole; 23. Bolt; 24. Wheel seat; 25. Moving wheel. Detailed Implementation

[0037] The present invention will now be described in further detail with reference to the accompanying drawings and specific embodiments.

[0038] like Figure 1-11 As shown, a near-field and far-field collaborative top-cutting method includes the following steps:

[0039] S1: Use single hydraulic props in conjunction with metal articulated top beams and cross beams to reinforce and support the roof of the roadway in advance. The reinforcement of the roadway in advance should be carried out before the working face is mined. Considering that it is better to make cracks outside the advance pressure range of the directional pre-splitting blast holes, the length of the reinforcement of the roadway in advance should be appropriately increased to 20-25m. That is, the length of the reinforcement of the roadway outside the coal wall of the working face is 20-25m. As the working face is mined, the construction is continuously extended outward, keeping the minimum length not less than 20m.

[0040] Generally, when the roof of a coal face is extremely thick, hard, and intact, the support for the roadway is very simple. Some coal mines even do not support the roof at all, using only anchor mesh support for the roadway sides during excavation. During mining, two rows of single hydraulic props in the advance support section are sufficient to meet the production and safety requirements. However, subsequent advance directional blasting will damage the integrity of the roadway. Relying solely on the traditional method of installing two rows of single hydraulic props may not be enough to maintain the stability of the roof after blasting, leading to roof collapse and ballast drop accidents, threatening the safety of workers in the roadway. Therefore, it is necessary to strengthen the support of the roadway in the traditional advance support section. The traditional method of support is simple but the support force is insufficient, so it needs to be strengthened. After strengthening the support, the advance support section is generally supported by 2-3 rows of anchor cables combined with 2-3 rows of single hydraulic point pillars and metal articulated roof beams, or 3-4 rows of single hydraulic props combined with metal articulated roof beams.

[0041] Taking a mining roadway with a uniform thickness of 16m diorite porphyry as the immediate roof, a width of 4.5m, and a height of 2.5m as an example, the coal face 20-25m ahead is reinforced with single hydraulic props in conjunction with metal articulated roof beams and cross beams. Single hydraulic props are installed at 1000mm above the upper side, 500mm below the lower side, and 1500mm below the designed roof cutting line, referred to as the upper, middle, and lower routes, with prop spacing of 500mm, 500mm, and 1000mm respectively. The support method is as follows: metal articulated roof beams and cross beams are used in conjunction with single props. The single hydraulic props for the upper and middle routes are installed under the cross beams, and the two cross beams are connected by articulated roof beams. The lower route uses single hydraulic props and articulated roof beams for support along the roadway. The beams must be articulated, and the prop heads must be securely fastened with wire. The articulation rate of the roof beams is ≥90%.

[0042] S2: Construct roof-cutting blast holes in the roadway roof. The blast holes are arranged in a row along the roadway direction, with the openings of all holes on a straight line. The holes are driven towards the coal face of the working face. The drilling angle of the blast holes is 10-15° with the vertical line, the depth is 7-15m, the diameter is 50-70mm, the hole spacing is 0.5-1m, the charge per hole is 4-10kg, and the construction position is 2.5-5m away from the coal face of the roadway.

[0043] The spacing between the boreholes is determined based on the specific rock hardness and the amount of explosive charge; the harder the rock, the more explosive charge is required.

[0044] S3: In roadways where roadway retention is required, move the support mechanism to a range of 10-25m from the coal face. This range of 10-25m from the coal face is not only the range of advanced directional pre-splitting, but also the range of advanced support pressure of the working face. The roadway roof is supported and held in place by the support mechanism. The support position is the end of the roof that is close to the working face.

[0045] S4: In roadways where roadway retention is required, advance directional pre-splitting blasting is carried out within a range of 18-25m from the coal wall. Before construction, it is confirmed that the reinforcement quality of the advance roadway is intact. Mining-grade Class III emulsion explosive is placed in the shaped charge tube and connected to the detonation device. The shaped charge tube with charge is installed in the completed blasting borehole. The detonator leads are straightened. Then the borehole is sealed, the wire is connected, and the blasting is initiated to carry out directional pre-splitting, so that cracks appear in the roof along the edge of the designed roadway retention. Then the blasting effect is checked.

[0046] S5: When the roof is a thick, hard rock layer, even if simple roof cutting and jointing can cut through the roof, it is not easy for the roof to collapse in the goaf. In this case, forced roof caving is required. Bidirectional forced roof caving is adopted, that is, forced roof caving blast holes are drilled into the roof of the working face from both roadways. Each roadway is set up with 3 sets of holes in a cycle. The 3 sets of holes are numbered A, B and C respectively. They are all oriented towards the goaf and form angles of 70°, 45° and 10° with the roadway in the plane, that is, 20°, 45° and 80° with the vertical lines of the two roadways. Group A is drilled with 3 holes, numbered A1, A2 and A2. Group A3 has angles of 10°, 15°, and 20° with the roof, respectively. Group B has two holes, numbered B1 and B2, with angles of 23° and 35° with the roof, respectively. Group C holes are set differently depending on the destination of the goaf retaining roadway. For roadways without goaf retaining roadway, Group C holes are set as two holes, numbered C1 and C2, with angles of 20° and 30° with the roof, respectively. For roadways with goaf retaining roadway, Group C holes are set as three holes, numbered C1, C2, and C3, with angles of 20°, 30°, and 40° with the roof, respectively.

[0047] Taking the example of a 16m thick diorite porphyry as the direct roof, and a 70m long coal face as an example, a two-way forced roof caving method is adopted, that is, forced roof caving blast holes are drilled into the roof of the working face from both roadways. Each roadway is set with 3 sets of holes in a cycle, with a cycle distance of 13m and a hole diameter of 50-70mm. The 3 sets of holes are numbered A, B, and C respectively. They are all oriented towards the goaf and form angles of 70°, 45°, and 10° with the roadway in the plane, that is, 20°, 45°, and 80° with the vertical lines of the two roadways. Group A has 3 holes, each with a depth of 38m, numbered A1, A2, A3, A4, A5, A6, A7, A8, A9, A1, A9 ... 2. A3, with angles to the roof of 10°, 15°, and 20° respectively. Group B drills two holes, each 18m deep, numbered B1 and B2, with angles to the roof of 23° and 35° respectively. Group C drills holes 17m deep. The holes are set differently depending on the destination of the goaf retaining roadway: for roadways without goaf retaining roadway, Group C holes are set as two holes, numbered C1 and C2, with angles to the roof of 20° and 30° respectively; for roadways with goaf retaining roadway, Group C holes are set as three holes, numbered C1, C2, and C3, with angles to the roof of 20°, 30°, and 40° respectively.

[0048] S6: Continuously load explosives into holes in groups A, B, and C, leaving 2-3m of the hole opening unloaded, then seal the hole, connect the wires, and detonate. Forced roof collapse blasting causes the roof of the goaf to collapse, increasing the small-angle blasting along the roadway direction and the goaf, i.e., densifying the holes in group C. Using the pre-splitting blasting cut as a free surface, fully blast and break the rock mass between the holes in group C and the pre-splitting cut, inducing it to collapse and accumulate to form a gangue belt, thus playing a supporting role in the process of the overlying strata collapsing or rotating and sinking.

[0049] S7: The technical parameters for forced roof caving need to be adjusted according to the specific geological conditions of the coal mining face. The main parameters for adjustment are the hardness of the rock strata, the integrity of the rock strata, and the minimum resistance line radius of the blasting. When the hardness and integrity of the roof rock strata are high and the minimum resistance line radius of the blasting is small, the included angle between holes in the same group should be reduced and the number of holes should be increased appropriately. Conversely, the included angle between holes should be increased and the number of holes should be reduced appropriately.

[0050] In particular, the small angle between the C-group borehole and the roadway is the key to the formation of gangue piles and the relief of the roadway roof collapse caused by blasting.

[0051] In the above methods and steps, Figure 1 , Figure 2 Schematic diagram for strengthening support of advanced roadways. Figure 3 , Figure 4 This is a schematic diagram of forced jacking for three groups of holes: Group A, Group B, and Group C. Figure 5 This is a diagram showing the situation before the forced caving. Figure 6 This is a diagram showing the forced caving out.

[0052] The above method first involves directional pre-splitting through directional pre-splitting blast holes constructed along the roadway direction, causing cracks to appear in the roof along the edge of the designed roadway. Then, forced roof caving blasting causes the roof of the goaf to collapse. Based on the traditional forced roof caving, small-angle blasting along the roadway direction and the goaf is added, i.e., densification of C-group holes. The pre-splitting blasting cuts are used as free surfaces to fully blast and break up the rock mass between the C-group holes and the pre-splitting cuts, inducing it to collapse and accumulate to form a gangue belt. This belt plays a supporting role during the collapse or rotation and subsidence of the overlying strata. In other words, through pre-splitting roof cutting and the improved forced roof caving in the goaf, the transmission of dynamic pressure from the goaf to the roadway area is blocked, and the roof of the roadway on the goaf side collapses to support the overlying rock mass, relieving the pressure on the roadway roof and ensuring the stability of the roadway roof along the goaf.

[0053] The support mechanism in S3 includes an L-shaped support plate 6; multiple connecting frames 7 are evenly distributed at the lower end of the support plate 6; a support block 8 is installed on the support plate 6 between two adjacent connecting frames 7; a support device is provided at the lower end of each connecting frame 7; the support device includes an A electric push rod 9; the output end of the upper end of the A electric push rod 9 is rotatably connected to a connector 10 via a rotating shaft, and the connector 10 and the connecting frame 7 are engaged; an inclined B electric push rod 11 is fixedly connected to the front side of the A electric push rod 9; the A electric push rod 9 and the B electric push rod 11... Pressure sensors 12 are fixedly connected to the lower ends of each pressure sensor 12, and the display screens 13 of the pressure sensors 12 are all mounted on the corresponding A electric push rod 9 and B electric push rod 11; a fixed base 14 is fixedly connected to the lower end of each pressure sensor 12; a fixed tooth 15 is fixedly connected to the lower end of each fixed base 14; a drive bar 16 is slidably connected to both ends of the A electric push rod 9; the upper end of each drive bar 16 is fixedly connected to the connector 10; a moving device is fixedly connected to the lower end of each drive bar 16; the connecting frame 7 is an L-shaped connecting frame 7, and the connecting frame 7 and The support plate 6 is fixedly connected; the lower end of the connecting frame 7 has a connecting port 17 that engages with the connecting head 10; the lower end of each support block 8 has a connecting groove 18 that engages with the single hydraulic prop; the support block 8 is divided into support block A 19 and support block B 20; the upper end face of support block A 19 is located at the upper inner side of the support plate 6; the upper end face and the rear end face of support block B 20 engage with the inner side of the support plate 6 at an angle, and the connecting groove 18 of support block B 20 is inclined; between two adjacent connecting frames 7, there are two mounting holes 21 on the support plate 6; Two mounting holes 21 correspond to support block A 19 and support block B 20 respectively; in the connecting groove 18, each support block 8 has a threaded hole 22 corresponding to the mounting hole 21; the mounting hole 21 and the threaded hole 22 are connected by bolts 23; the lower end of the threaded hole 22 is higher than the lower end of the bolt 23; the moving device includes a semi-circular tubular wheel seat 24; the upper end of the wheel seat 24 is fixedly connected to the drive bar 16, and the wheel seat 24 corresponds to support block A 19; each wheel seat 24 has a fixedly connected movable wheel 25 at its lower end; the movable wheel 25 is a self-locking universal wheel;

[0054] The remaining coal seam refers to the solid coal seam in the roadway; the detonation device can be a mine detonator or detonating cord; the connecting groove 18 corresponds to the upper connector 10 of the single hydraulic support; the support block 8 is divided into two structures, one is support block A 19, and the other is support block B 20, which facilitates auxiliary support for vertical and inclined applications, as shown in the specific structure. Figure 11 As shown;

[0055] When the support mechanism is in operation, the support plate 6 is moved to the position where advanced directional pre-splitting is required, i.e., within a range of 18-25m from the coal face. The L-shape of the support plate 6 fits the roof and sidewall together. The support position is the end of the roof plate closest to the working face. Then, the A electric push rod 9 is activated. The A electric push rod 9 drives the connector 10 on it to move. The connector 10 moves and engages with the connector port 17 on the connecting frame 7. The A electric push rod 9 drives the upper end of the support plate 6 to fit against the roof. The lower end of the A electric push rod 9 passes through... The fixing teeth 15 on the fixed base 14 are connected to the bottom surface of the roadway. The installation of electric push rod 9 and support plate 6 strengthens the support of the pre-cracked end of the roof. When the pre-cracked rock is cut, depending on the location of the pre-crack, the pre-cracked rock will cause a large weight pressure on the end of the roof that is close to the working face. This can easily cause serious deformation of the roadway and result in significant losses and impacts. However, by strengthening the support with electric push rod 9 and support plate 6, the support effect is better, preventing excessive deformation of the roadway from causing an impact.

[0056] When electric actuator A 9 supports the support plate 6, electric actuator B 11 is activated. Electric actuator B 11 extends and is fixed in the roadway by the fixing teeth 15 on the fixing seat, supporting one side of electric actuator A 9. Because the pre-split rocks will cause a large weight pressure on the end of the roof near the working face, the inclined electric actuator B 11 provides better force distribution for electric actuator A 9 and makes it easier to strengthen the support.

[0057] Pressure sensors 12 installed at the lower ends of electric actuators A and B can detect the pressure values ​​on electric actuators A and B. The pressure values ​​can be easily viewed on the display screen 13. When the pressure value at electric actuator A is too high, a single hydraulic support is added. Each support block 8 has a connecting groove 18 at its lower end that mates with the single hydraulic support. The single hydraulic support connects to the support block 8 on the support plate 6, and the single hydraulic support is engaged in the connecting groove 18, thus holding the support plate 6 in place. The single hydraulic support replaces the A... The electric push rod 9 supports the support plate 6, strengthens the support of the pre-splitting roadway, and prevents the roadway deformation from causing an impact. When the pressure sensor 12 on the electric push rod 11 detects that the pressure value is too high, it means that the support plate 6 is under great pressure from the tilting force. The single hydraulic prop and the connecting groove 18 on the tilted support block 20 can be matched. The single hydraulic prop will then tilt to support the tilt angle of the support plate 6 through the support block 20, resulting in better force support. The tilted single hydraulic prop can better support the tilting force on the support plate 6.

[0058] When pressure sensor 12 detects excessive pressure, it uses individual hydraulic supports to prevent excessive deformation of the tunnel. After the individual hydraulic supports are in place, all A electric push rods 9 are retracted. The retraction of A electric push rods 9 causes the connector 10 and the drive bar 16 to move downwards. The drive bar 16 then causes the wheel seat 24 and the moving wheel 25 to move downwards. After the moving wheel 25 contacts the ground, A electric push rods 9 continue to retract. The fixed tooth 15 below A electric push rods 9 will move upwards and no longer contact the ground. At this point, it will... Figure 9 As shown, only the moving wheel 25 is in contact with the ground, which makes it easy for the support device to move to the next pre-crack area. Then, the above operations are performed by placing a new support plate 6 to strengthen the support and to detect the weight pressure value after pre-cracking. This facilitates movement and can be used in conjunction with advanced pre-cracking for safer auxiliary support. When the weight pressure value after pre-cracking is detected to be small, the electric push rod A 9 can be retracted directly, and then it can be used for the next pre-crack area.

[0059] The semi-circular tubular wheel seat 24 is more convenient for use when reinforced by a single hydraulic prop. The threaded holes 22 on the A support block 19 and the B support block 20, along with the corresponding mounting holes 21 on the support plate 6, make it easy to install either the A support block 19 or the B support block 20 onto the support plate 6 for auxiliary reinforcement, making it more convenient to use. The connection of the bolts 23 does not affect the fit between the single hydraulic prop and the connecting groove 18 when the bolts 23 are connected to the A support block 19 or the B support block 20.

[0060] It should be understood that the specific embodiments described above are merely illustrative or explanatory of the principles of the invention and do not constitute a limitation thereof. Therefore, any modifications, equivalent substitutions, improvements, etc., made without departing from the spirit and scope of the invention should be included within the protection scope of the invention. Furthermore, the appended claims are intended to cover all variations and modifications falling within the scope and boundaries of the appended claims, or equivalent forms of such scope and boundaries.

Claims

1. A method for coordinated near-field and far-field top cutting, characterized in that: Includes the following steps: S1: Use single hydraulic props in conjunction with metal articulated top beams and cross beams to reinforce and support the roof of the roadway in advance. The reinforcement of the roadway in advance should be carried out before the working face is mined. Considering that it is better to make cracks outside the advance pressure range of the directional pre-splitting blast holes, the length of the reinforcement of the roadway in advance should be appropriately increased to 20-25m. That is, the length of the reinforcement of the roadway outside the coal wall of the working face is 20-25m. As the working face is mined, the construction is continuously extended outward, keeping the minimum length not less than 20m. S2: Construct roof-cutting blast holes in the roadway roof. The blast holes are arranged in a row along the roadway direction, with the openings of all holes on a straight line. The holes are driven towards the coal face of the working face. The drilling angle of the blast holes is 10-15° with the vertical line, the depth is 7-15m, the diameter is 50-70mm, the hole spacing is 0.5-1m, the charge per hole is 4-10kg, and the construction position is 2.5-5m away from the coal face of the roadway. S3: In roadways where roadway retention is required, move the support mechanism to a range of 10-25m from the coal face. This range of 10-25m from the coal face is not only the range of advanced directional pre-splitting, but also the range of advanced support pressure of the working face. The roadway roof is supported and held in place by the support mechanism. The support position is the end of the roof that is close to the working face. S4: In roadways where roadway retention is required, advance directional pre-splitting blasting is carried out within a range of 18-25m from the coal wall. Before construction, it is confirmed that the reinforcement quality of the advance roadway is intact. Mining-grade Class III emulsion explosive is placed in the shaped charge tube and connected to the detonation device. The shaped charge tube with charge is installed in the completed blasting borehole. The detonator leads are straightened. Then the borehole is sealed, the wire is connected, and the blasting is initiated to carry out directional pre-splitting, so that cracks appear in the roof along the edge of the designed roadway retention. Then the blasting effect is checked. S5: When the roof is a thick, hard rock layer, even if simple roof cutting and jointing can cut through the roof, it is not easy for the roof to collapse in the goaf. In this case, forced roof caving is required. Bidirectional forced roof caving is adopted, that is, forced roof caving blast holes are drilled into the roof of the working face from both roadways. Each roadway is set up with 3 sets of holes in a cycle. The 3 sets of holes are numbered A, B and C respectively. They are all oriented towards the goaf and form angles of 70°, 45° and 10° with the roadway in the plane, that is, 20°, 45° and 80° with the vertical lines of the two roadways. Group A is drilled with 3 holes, numbered A1, A2 and A2. Group A3 has angles of 10°, 15°, and 20° with the roof, respectively. Group B has two holes, numbered B1 and B2, with angles of 23° and 35° with the roof, respectively. Group C holes are set differently depending on the destination of the goaf retaining roadway. For roadways without goaf retaining roadway, Group C holes are set as two holes, numbered C1 and C2, with angles of 20° and 30° with the roof, respectively. For roadways with goaf retaining roadway, Group C holes are set as three holes, numbered C1, C2, and C3, with angles of 20°, 30°, and 40° with the roof, respectively. S6: Continuously charge explosives in holes of groups A, B, and C, leaving 2-3m uncharged at the hole opening. Then seal the holes, connect the wires, and detonate. Forced roof caving blasting causes the roof of the goaf to collapse, increasing the small-angle blasting along the roadway direction and the goaf, i.e., densifying the holes of group C. Utilize the pre-splitting blasting cut as a free surface to fully blast and break up the rock mass between the holes of group C and the pre-splitting cut, inducing it to collapse and accumulate to form a gangue belt, thus playing a supporting role during the collapse or subsidence of the overlying strata. Charge, seal, and connect the explosives just before the forced roof caving blast holes enter the goaf. Blasting is carried out as soon as they enter the goaf. Before blasting, strengthen the support within 20m of the blasting section to prevent blasting vibration from damaging some roadways that need maintenance. After blasting, check the blasting effect of forced roof caving to see if the goaf has collapsed and whether the roof of the remaining roadway is stable. S7: The technical parameters for forced roof caving need to be adjusted according to the specific geological conditions of the coal mining face. The parameters for adjustment are the rock layer hardness, rock layer integrity, and minimum resistance line radius of blasting. When the roof rock layer has high hardness and good integrity, and the minimum resistance line radius of blasting is small, the included angle between holes in the same group should be reduced and the number of holes should be increased appropriately. Conversely, the included angle between holes should be increased and the number of holes should be reduced appropriately. The support mechanism in S3 includes an L-shaped support plate (6); multiple connecting frames (7) are evenly distributed at the lower end of the support plate (6); between two adjacent connecting frames (7), a support block (8) is installed on the support plate (6); and a support device is provided at the lower end of the connecting frame (7). The support device includes an electric push rod A (9); the output end of the upper end of the electric push rod A (9) is rotatably connected to a connector (10) via a rotating shaft, and the connector (10) and the connecting frame (7) are engaged; an inclined electric push rod B (11) is fixedly connected to the front side of the electric push rod A (9); a pressure sensor (12) is fixedly connected to the lower end of both the electric push rod A (9) and the electric push rod B (11), and the display screen (13) of the pressure sensor (12) is set on the corresponding electric push rod A (9) and electric push rod B (11); a fixed base (14) is fixedly connected to the lower end of the pressure sensor (12); a fixed tooth (15) is fixedly connected to the lower end of the fixed base (14); a drive bar (16) is slidably connected to both the left and right ends of the electric push rod A (9); the upper end of the drive bar (16) is fixedly connected to the connector (10); a moving device is fixedly connected to the lower end of the drive bar (16). The connecting frame (7) is an L-shaped connecting frame (7), and the connecting frame (7) is fixedly connected to the support plate (6); the lower end of the connecting frame (7) has a connecting port (17) that engages with the connecting head (10); Each support block (8) has a connecting groove (18) at its lower end that matches the single hydraulic prop (1).

2. The near-field and far-field coordinated top-cutting method according to claim 1, characterized in that: The support block (8) is divided into support block A (19) and support block B (20); the upper end face of support block A (19) is located at the upper end of the inner side of support plate (6); the upper end face and rear end face of support block B (20) are fitted with the inner side of support plate (6) at the angle, and the connecting groove (18) of support block B (20) is inclined.

3. The near-field and far-field coordinated top-cutting method according to claim 2, characterized in that: Between two adjacent connecting frames (7), there are two mounting holes (21) on the support plate (6); the two mounting holes (21) correspond to support block A (19) and support block B (20) respectively.

4. The near-field and far-field coordinated top-cutting method according to claim 3, characterized in that: Inside the connecting groove (18), the support block (8) has a threaded hole (22) corresponding to the mounting hole (21); the mounting hole (21) and the threaded hole (22) are connected by a bolt (23); the lower end of the threaded hole (22) is higher than the lower end of the bolt (23).

5. The near-field and far-field coordinated top-cutting method according to claim 2, characterized in that: The moving device includes a semi-circular tubular wheel seat (24); the upper end of the wheel seat (24) is fixedly connected to the drive bar (16), and the wheel seat (24) is correspondingly matched with the A support block (19); the lower end of the wheel seat (24) is fixedly connected with a moving wheel (25).

6. The near-field and far-field coordinated top-cutting method according to claim 1, characterized in that: The caster wheel (25) is a self-locking universal wheel.

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