Method for blasting roof cutting and pressure releasing of overburden strata of coal pillar lateral goaf of top coal caving face

Abstract

The application discloses a kind of overburden strata blasting top cutting pressure relief method in coal pillar lateral goaf of top coal caving face, to effectively solve the problem of poor blasting effect in hard rock strata overburden of coal pillar lateral goaf in prior art.To this end, the overburden strata blasting top cutting pressure relief method in coal pillar lateral goaf of top coal caving face provided by the application drills inclined upward boreholes in coal seam roof towards goaf in coal pillar roadway, and makes the boreholes penetrate through hard rock strata of coal seam roof.The boreholes are arranged in front of stress concentration zone of coal seam roof and distributed from dense to sparse along the advancing direction of coal mining face, then uncoupled charging is carried out in each borehole, and detonation is started from borehole dense area, to carry out pre-splitting of coal seam roof above goaf.After that, with the advancing of coal mining face, stress concentration zone of coal seam roof moves forward continuously, and further promotes the further collapse of coal seam roof above goaf, to realize overburden strata blasting top cutting pressure relief in goaf.

Description

Technical Field

[0001] This invention belongs to the field of coal seam mining technology, and in particular relates to a method for blasting and cutting the overlying strata of the coal pillar in the goaf of a top coal caving working face to relieve pressure. Background Technology

[0002] The high-stress triangle caused by the overhanging roof on one side of the coal pillar goaf is a major cause of the restriction on the retention of small coal pillars in the working face, leading to large deformation of the mining roadway and even rockbursts. To reduce or even eliminate the impact of the triangle, the current main method is roof hydraulic fracturing with water injection to change the continuity of the roof, promote the removal of the roof on one side of the coal pillar goaf, eliminate the roof triangle, and relieve pressure on the coal pillar and the mining roadway. This method is only suitable for hydraulic fracturing of the roof in thin and medium-thick coal seams when mining the full height in one go. However, for top coal caving, due to the large-scale seepage of water along the coal-rock interface during the water injection process, it is easy to cause large-scale delamination of the top coal, leading to large-scale collapse of the top coal in the roadway, resulting in mass casualties and significant property damage.

[0003] To address the issue of roof collapse on one side of the goaf in coal pillar mining, the current effective method is drill-and-blast. This method has a short working time and fast construction speed. However, the current blasting method involves non-coupled and uniform charge in the boreholes. During detonation, all boreholes are detonated simultaneously. Since the thick and hard roof is almost an integral structure on the goaf side of the coal pillar, the drilling and blasting is subject to significant rock mass clamping. Moreover, the free surface of the initial borehole is only the roof fracture surface on the goaf side, which undoubtedly severely restricts the effectiveness of the blasting. Summary of the Invention

[0004] The main objective of this invention is to provide a method for blasting and cutting the roof to relieve pressure in the overlying strata of the coal pillar side goaf in a top coal caving face, which aims to effectively solve the problem of poor blasting effect of hard overlying strata in the coal pillar side goaf in the prior art.

[0005] To address this, the present invention provides a method for blasting and cutting the overlying strata of the coal pillar in a top-coal caving face to relieve pressure in the goaf. This method involves drilling upward-sloping boreholes in the coal pillar roadway towards the goaf on the coal seam roof, penetrating the hard strata of the coal seam roof. These boreholes are arranged in front of the pre-stress concentration zone of the coal seam roof and distributed from dense to sparse along the direction of the coal face advance. Decoupled charges are then placed in each borehole, with detonation starting from the densely packed borehole area to pre-fracture the coal seam roof above the goaf. As the coal face advances, the pre-stress concentration zone of the coal seam roof continuously moves forward, further causing the coal seam roof above the goaf to collapse, thereby achieving blasting and cutting the overlying strata of the goaf to relieve pressure.

[0006] Specifically, the explosive charge structure inside the borehole includes a shaped charge placed in the hard rock strata and a regular charge placed in the soft rock strata, and the outer end of the borehole is sealed with a plug.

[0007] Specifically, the shaped charge includes a flexible metal mesh cylinder, granular explosives filled inside the flexible metal mesh cylinder, and a spreading assembly that causes the flexible metal mesh cylinder to contract axially and expand radially to abut against the wall of the borehole.

[0008] Specifically, the spreading component includes a telescopic rod disposed inside the flexible metal mesh cylinder. The two ends of the telescopic rod are fixedly connected to the two ends of the flexible metal mesh cylinder, respectively. The telescopic rod is also provided with a support frame that spreads the two ends of the flexible metal mesh cylinder. The telescopic rod includes an upper tube section and a lower tube section. The upper end of the lower tube section is slidably fitted onto the upper tube section. One end of the support frame is hinged to the lower tube section, and the other end is hinged to the flexible metal mesh cylinder. The upper tube section is provided with a spring pin, and the inner wall of the lower tube section is provided with a pin hole that cooperates with the spring pin.

[0009] During loading, a push rod is used to send the shaped charge into the borehole until it comes into contact with the ordinary charge cartridges loaded in the soft rock layer above. Then, the push rod continues to push the shaped charge, causing the lower pipe section to move upward relative to the upper pipe section. This forces the support frame to open the flexible metal mesh cylinder and continuously squeeze and push the granular explosives inside the flexible metal mesh cylinder, so that the flexible metal mesh cylinder abuts against the wall of the borehole until the spring pin is engaged in the pin hole.

[0010] Specifically, a number of pull ropes are also connected between the telescopic rod and the flexible metal mesh cylinder to detach the side portion of the flexible metal mesh cylinder from the borehole wall.

[0011] Specifically, the telescopic rod is a tube made of iron filings, and the tube is filled with single-element high explosive. When detonated, the high explosive has the characteristics of fast detonation speed and strong detonation capability, which detonates the surrounding particulate explosives almost simultaneously. At the same time, the iron filings generated by the breaking of the tube act as sensitizers, activating the surrounding explosive particles over a wide range.

[0012] Specifically, the borehole is equipped with staged energy reduction ends at both ends of the charging structure. One staged energy reduction end is located at the top of the borehole, and the other staged energy reduction end is located between the charging structure and the plug, and above the coal-rock interface.

[0013] Specifically, the graded energy reduction end includes two natural rubber substrates arranged vertically and connected by a connecting rod, and multiple metal material layers spaced apart on the connecting rod. Each metal material layer has an outwardly convex conical surface facing the outer surface of the charge structure.

[0014] Specifically, a blank reserved hole is set between any two adjacent boreholes, where no explosive charge will be placed.

[0015] Specifically, the amount of explosive charge per unit length of borehole is determined based on the lithology of the rock strata, the spacing between adjacent boreholes, and the properties of the explosive. The diameter and density of the explosive are then determined based on the amount of explosive charge per unit length.

[0016] Specifically, the borehole inclination angle θ is determined by combining the thickness a of the coal seam roof and the width b of the coal pillar, tanθ=a / b.

[0017] Compared with the prior art, the present invention has the following beneficial effects:

[0018] 1. In the coal pillar roadway, drill inclined upward boreholes towards the goaf on the coal seam roof. Decoupled charges are used in the boreholes to blast the coal seam roof. After the blasting of the coal seam roof in the goaf, significant large cracks are formed along the direction of the borehole arrangement, thus creating a new blasting free surface and promoting the rapid collapse of the roof. Moreover, by arranging the boreholes in front of the pre-stress concentration zone of the coal seam roof and distributing them from dense to sparse along the direction of the coal mining face advance, the existence of pre-stress concentration zones in the roof strata of the coal mining face can be fully utilized. The fact that these pre-stress concentration zones move forward with the advancement of the working face means that as the coal mining face continues to advance, the huge roof pressure will cause the blasted coal seam roof above the goaf to collapse further, thereby achieving the blasting and pressure relief of the overlying strata in the goaf.

[0019] 2. The blast holes are arranged in a dense-to-sparse distribution, and blank reserved holes are set on the adjacent sides when the explosive charge is detonated, which is a form of intermittent detonation. This arrangement of detonation can increase the propagation distance of the shock wave through the reserved holes, increase the extension of the crack, and make the upper hard top plate collapse well, thereby improving the blasting effect.

[0020] 3. For sections of hard rock strata, use shaped charge explosives with elastic columnar filling of loose particles. Place the charge longer than the section of hard rock strata and push it to compress and thicken it, which indirectly increases the charge amount. This allows the charge to fit tightly against the inner wall of the borehole, greatly reducing the coupling distance and thus achieving precise blasting of the borehole. This solves the problem of poor blasting effect in hard rock strata due to insufficient charge amount.

[0021] 4. By connecting the telescopic rod and the flexible metal mesh cylinder with a rope, the side wall of the flexible metal mesh cylinder is forced to detach from the borehole wall, thereby forming a condensed energy cavity. This causes the detonation products to form a condensed energy flow with high pressure and velocity, which has a strong impact on the borehole wall and improves the overall blasting effect.

[0022] 5. The staged energy reduction end set at the top of the borehole can buffer and absorb shock waves, preventing them from damaging the aquifer in the overlying strata and causing water inrush accidents in the roadway. The staged energy reduction end set at the junction of the coal seam and the overlying strata can realize multi-level reflection and energy reduction of stress waves, protect the coal-rock interface, and minimize the impact of stress waves on coal-rock separation. Attached Figure Description

[0023] To more clearly illustrate the technical solutions in the embodiments of the present invention, the accompanying drawings used in the description of the embodiments will be briefly introduced below. Obviously, the accompanying drawings described below are only some embodiments of the present invention. For those skilled in the art, other drawings can be obtained based on these drawings without creative effort.

[0024] Figure 1 This is a cross-sectional view of a coal seam according to an embodiment of the present invention;

[0025] Figure 2 This is a schematic diagram of the borehole distribution according to an embodiment of the present invention;

[0026] Figure 3 This is a schematic diagram of a large crack formed after drilling and blasting, according to an embodiment of the present invention.

[0027] Figure 4 This is a schematic diagram of the collapse of the coal seam roof in a goaf area according to an embodiment of the present invention;

[0028] Figure 5 This is a schematic diagram of the borehole charging structure according to an embodiment of the present invention;

[0029] Figure 6 This is a schematic diagram of the shaped charge explosive package before compression, according to an embodiment of the present invention;

[0030] Figure 7 This is a schematic diagram of a compressed shaped charge according to an embodiment of the present invention;

[0031] Figure 8 yes Figure 7 Enlarged view of point A in the middle;

[0032] Figure 9 This is a schematic diagram of the graded energy reduction end structure involved in an embodiment of the present invention;

[0033] Figure 10 This is a schematic diagram of the distribution of blank reserved holes according to an embodiment of the present invention;

[0034] The components include: 1. Coal pillar roadway; 2. Goaf; 3. Coal seam roof; 4. Borehole; 5. Hard rock strata; 6. Coal mining face; 7. Aquifer; 8. shaped charge; 801. Flexible metal mesh cylinder; 802. Particle explosive; 803. Telescopic rod; 8031. Upper pipe section; 8032. Lower pipe section; 804. Support frame; 8041. Connecting rod; 805. Spring pin; 806. Pin hole; 807. Pull rope; 9. Ordinary explosive cartridge; 10. Graded energy reduction end; 101. Natural rubber matrix; 102. Metal material layer; 11. Weak rock strata; 12. Shaped charge cavitation; 13. Single-element high explosive; 14. Blank reserved hole; 15. Large fracture; 16. Plug. Detailed Implementation

[0035] The technical solutions of the embodiments of the present invention will be clearly and completely described below with reference to the accompanying drawings. Obviously, the described embodiments are only some embodiments of the present invention, and not all embodiments. Based on the embodiments of the present invention, all other embodiments obtained by those skilled in the art without creative effort are within the scope of protection of the present invention.

[0036] 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.

[0037] 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 technical features indicated. Thus, a feature defined as "first" or "second" may explicitly or implicitly include one or more of that feature. In the description of this invention, "a plurality of" means two or more, unless otherwise explicitly specified.

[0038] See Figures 1-4A method for blasting and cutting the roof to relieve pressure in the overlying strata of a coal pillar in a top-coal caving face involves drilling inclined boreholes 4 in the coal pillar roadway 1 towards the goaf 2 on the coal seam roof 3, with the boreholes 4 penetrating the hard rock strata 5 of the coal seam roof 3. The boreholes 4 are arranged in front of the advanced stress concentration zone of the coal seam roof 3 and distributed from dense to sparse along the advancing direction of the coal mining face 6. The distribution from dense to sparse means that the spacing between the boreholes 4 is continuously increasing. Then, decoupled charges are carried out in each borehole 4, and detonation begins in the dense area of ​​the boreholes 4. After the blasting is completed, the advanced stress concentration zone of the coal seam roof 3 moves forward continuously as the coal mining face 6 advances, causing the blasted coal seam roof 3 above the goaf 2 to collapse further, thereby achieving blasting and cutting the roof to relieve pressure in the overlying strata of the goaf 2.

[0039] In this embodiment, inclined boreholes 4 are drilled in the coal pillar roadway 1 towards the goaf 2 and onto the coal seam roof 3. Decoupled charges are placed in the boreholes 4 to blast the coal seam roof 3. After the blasting of the coal seam roof 3 in the goaf 2, significant large cracks 15 are formed along the direction of the boreholes 4, thereby forming a new blasting free surface and promoting the rapid collapse of the roof. Moreover, by arranging the boreholes 4 in front of the pre-stress concentration zone of the coal seam roof 3 and distributing them from dense to sparse along the advancing direction of the coal mining face 6, the pre-stress concentration zone of the roof strata of the coal mining face 6 can be fully utilized. This zone moves forward with the advancement of the working face, so that as the coal mining face 6 continues to advance, the huge roof pressure will cause the blasted coal seam roof 3 above the goaf 2 to collapse further, thereby achieving the blasting and pressure relief of the overlying strata of the goaf 2.

[0040] See Figure 1 and Figure 5 In this embodiment, the charge structure inside the borehole 4 includes a shaped charge 8 placed in the hard rock layer 5 section and a regular charge 9 placed in the soft rock layer 11 section. The outer end of the borehole 4 is sealed by a plug 16.

[0041] The current drill-and-blast method uses conventional drill (blast) holes and conventional charge quantities for blasting pre-splitting. While it can achieve blasting fracturing effects in ordinary rock strata, the blasting fracturing effect is poor in hard rock strata 5 due to insufficient charge quantity. Usually, secondary blasting or an overall increase in charge quantity is required. Because the conventional advanced blast hole has insufficient charge space, it is usually necessary to use large-diameter drill holes 4 or enlarge the drill hole in the hard rock strata 5 to increase the charge quantity in order to achieve the expected pre-splitting effect in the hard rock strata 5. This will inevitably reduce the drilling speed of the drill hole 4, delay the advancement of the working face, and restrict the high-yield and efficient mining of the working face.

[0042] To address this, this application proposes a specially structured shaped charge 8. For the hard rock stratum 5 section, an elastic columnar shaped charge 8 filled with loose particles is used. The shaped charge 8 is placed longer than the hard rock stratum 5 section. During loading, a push rod is used to push it to compress and thicken it, which indirectly increases the amount of explosive. This allows the charge to fit tightly against the inner wall of the borehole 4 (blast hole), thereby achieving precise blasting of the borehole 4 and solving the problem of poor blasting effect due to insufficient explosive amount in the hard rock stratum 5.

[0043] See Figures 6-8 Specifically, the aforementioned shaped charge 8 includes a flexible metal mesh cylinder 801, granular explosive 802 filled in the flexible metal mesh cylinder 801, and a spreading component that causes the flexible metal mesh cylinder 801 to axially contract and radially expand so that it abuts against the wall of the borehole.

[0044] Specifically, the spreading component includes a telescopic rod 803 disposed inside the flexible metal mesh cylinder 801. The two ends of the telescopic rod 803 are fixedly connected to the two ends of the flexible metal mesh cylinder 801, respectively. The telescopic rod 803 is also provided with a support frame 804 that spreads the two ends of the flexible metal mesh cylinder 801. The telescopic rod 803 includes an upper tube section 8031 ​​and a lower tube section 8032. The upper end of the lower tube section 8032 is fitted and slidably sleeved outside the upper tube section 8031. One end of the support frame 804 is hinged to the lower tube section 8032, and the other end is hinged to the flexible metal mesh cylinder 801. The upper tube section 8031 ​​is provided with a spring pin 805, and the inner wall of the lower tube section 8032 is provided with a pin hole 806 that cooperates with the spring pin 805.

[0045] During loading, the push rod is used to send the shaped charge 8 into the borehole 4 until it contacts the ordinary charge 9 loaded at the weak rock layer 11 above. Then the push rod continues to push the bottom of the shaped charge 8. Due to the obstruction of the ordinary charge 9 above, the shaped charge 8 cannot move upward as a whole. At this time, driven by the push rod, the lower pipe section 8032 will move upward relative to the upper pipe section 8031, forcing the support frame 804 to open the flexible metal mesh cylinder 801 while continuously squeezing and pushing the granular explosive 802 inside the flexible metal mesh cylinder 801. This causes the flexible metal mesh cylinder 801 to expand radially and abut against the wall of the borehole 4 until the spring pin 805 is engaged in the pin hole 806, locking the lower pipe section 8032.

[0046] In this embodiment, the shaped charge 8 can achieve axial compression and radial expansion under the pushing action of the push rod. While compressing the granular explosive 802 inside the flexible metal mesh cylinder 801, it also makes the flexible metal mesh cylinder 801 abut against the wall of the borehole 4. By changing the thickness of the shaped charge 8, the charge per unit length and the detonation intensity of the borehole 4 can be changed, thereby achieving refined blasting of the borehole 4. This solves the problem of poor blasting effect in hard rock strata 5 due to insufficient charge in conventional borehole 4. Since conventional borehole 4 (blast hole) can be directly used for charging, efficient blasting of hard rock strata 5 can be achieved without enlarging the borehole 4, enabling high-yield and efficient mining of the working face.

[0047] See Figure 7 In some embodiments, a plurality of pull ropes 807 are uniformly connected between the telescopic rod 803 and the flexible metal mesh cylinder 801, which detach the side portion of the flexible metal mesh cylinder 801 from the borehole 4 wall. This design allows the flexible metal mesh cylinder 801 to partially detach from the borehole 4 wall when it is squeezed by the particulate explosive 802 and comes into contact with the borehole 4, thereby forming a shaped energy cavity 12. This causes the detonation products to form a shaped energy flow with high pressure and velocity, resulting in a powerful impact on the borehole wall and improving the overall blasting effect.

[0048] Understandably, in the actual design, one support frame 804 is set at the upper end and one at the lower end of the lower pipe section 8032. Each support frame 804 includes multiple connecting rods 8041 evenly arranged circumferentially along the lower pipe section 8032. One end of each connecting rod 8041 is hinged to the lower pipe section 8032, and the other end extends forward and outward and is hinged to the flexible metal mesh cylinder 801. In addition, the lower end face of the flexible metal mesh cylinder 801 is connected to the connecting rods 8041 of the lower support frame 804 by multiple cable ties. This design allows the lower end of the shaped charge 8 to be moved upward and compressed as a whole, thus not affecting the subsequent loading of ordinary explosive cartridges 9 into the weak rock layer 11.

[0049] It should be explained that, to prevent the loose granular explosive 802 from leaking out of the flexible metal mesh cylinder 801, during the expansion process of the flexible metal mesh cylinder 801, the mesh diameter of the flexible metal mesh cylinder 801 must always be smaller than the particle size of the granular explosive 802. The granular explosive 802 can be an ammonium nitrate explosive.

[0050] See Figure 6 and Figure 7In other embodiments, the telescopic rod 803 is a tube made of iron filings. The iron filings can be bonded or pre-sintered to form a tube with a certain strength. Then, the tube is filled with a single-element high explosive 13, such as TNT RDX. When detonated, the high explosive 13 has the characteristics of fast detonation speed and strong detonation ability, which detonates the surrounding particulate explosives 802 almost simultaneously, thereby significantly improving the blasting effect. At the same time, the iron filings generated by the breaking of the tube act as sensitizers, activating the surrounding explosive particles over a wide range.

[0051] The stress waves from drill-and-blast blasting can easily cause fracturing of the rock strata surrounding the top of borehole 4, potentially leading to flooding or other accidents as the aquifer 7 above the roof is exposed. Since the stress waves are compression waves, when propagating from a medium with higher impedance to a medium with lower impedance, they are easily reflected at the interface, causing part of the compression wave to transform into a tensile wave. The greater the difference in impedance between the two media, the stronger the interface reflection ability, and the greater the tensile stress generated. For coal and rock, their tensile strength is far less than their compressive strength. Therefore, when the stress waves propagate to the bottom of borehole 4, the coal-rock interface is more prone to tensile failure, leading to delamination at the coal-rock interface. To prevent localized disasters caused by delamination, the commonly used method is to expend significant manpower and resources on localized reinforcement support.

[0052] See Figure 5 In some embodiments, staged energy reduction ends (10) are provided at both ends of the charge structure inside the borehole. One staged energy reduction end (10) is located at the top of the borehole (4), and the other staged energy reduction end (10) is located between the charge structure and the plug (16), and is located 0.5-1m above the coal-rock interface.

[0053] In this embodiment, by setting a graded energy reduction end 10 at the top of the borehole 4, the shock wave can be buffered and absorbed to prevent it from damaging the aquifer 7 in the overlying rock layer and causing water inrush accident in the roadway. The graded energy reduction end 10 set at the junction of the coal seam and the overlying rock layer can realize multi-level reflection and energy reduction of stress wave, protect the coal-rock interface, and minimize the impact of stress wave on coal-rock separation.

[0054] See Figure 9 Specifically, the graded energy reduction end 10 includes two natural rubber substrates 101 arranged vertically and connected by a connecting rod, and multiple metal material layers 102 spaced apart on the connecting rod. Each metal material layer 102 has an outwardly convex conical surface facing the outer surface of the charge structure, and the multiple metal material layers 102 are connected in series by the connecting rod in the middle of the natural rubber substrates 101. In this embodiment, there are three metal material layers 102, namely an aluminum metal material layer 102, an iron metal material layer 102, and a tungsten metal material layer 102, and the conical surface is a conical surface with a cone angle of 45 degrees.

[0055] When loading explosives into borehole 4, firstly, a staged energy-reducing end cap 10 is installed at the top of borehole 4. Then, based on the rock strata properties, using a push rod, a shaped charge charge 8 is loaded into the hard rock stratum 5, and a regular explosive charge 9 is loaded into the weak rock stratum 11. After loading, a similar measure is taken to install the staged energy-reducing end cap 10 at the coal-rock interface. Finally, borehole 4 is sealed with a plug. Upon detonation, the shaped charge charge 8 and the regular explosive charge 9 inside borehole 4 detonate simultaneously. The staged energy-reducing end cap 10 utilizes the property of different wave impedances between different media. A large difference in wave impedance and a suitable incident angle will cause the stress wave generated by the explosion to be reflected and transmitted. The end is equipped with metal materials with different wave impedance values ​​as the medium. The wave impedance difference between the free surface of the air and the metal material is relatively large. At the same time, a 45° tilt is used to form a suitable incident angle, so that the stress wave generated by the explosion can be reflected to the side wall of the borehole, reducing the damage to the top and coal seam. The stress wave is reflected at the free surface of different metal materials (the wave impedance of aluminum, iron and tungsten metal materials also has a large difference). Even if some energy is transmitted into the interior and continues to propagate, multiple reflections can occur to form a reduction effect. Finally, natural rubber is provided as a buffer material to minimize the transmitted energy propagating to the top.

[0056] In this implementation, the graded energy reduction end 10 is cylindrical in shape and is installed in the borehole. The graded energy reduction end 10 set at the top of the borehole 4 can buffer and absorb the shock wave, preventing it from damaging the aquifer 7 in the overlying rock layer and causing seepage, thus preventing a water inrush accident in the roadway. The graded energy reduction end 10 set at the junction of the coal seam and the overlying rock layer can realize multi-level reflection and energy reduction of the stress wave, protect the coal-rock interface, and minimize the impact of the stress wave on the coal-rock separation.

[0057] See Figure 10 Specifically, a blank reserved hole 14 without charging is provided between any two adjacent drill holes 4. In this embodiment, the blast holes are arranged in a dense-to-sparse distribution and blank reserved holes 14 are provided on the adjacent side when the charging hole is detonated, which is a form of intermittent detonation. This arrangement of detonation can increase the propagation distance of the shock wave through the reserved hole, increase the extension of the crack, and make the upper hard top plate collapse well, thereby achieving the purpose of improving the blasting effect.

[0058] In this embodiment, since the thick, hard roof triangular area is approximately a single structure on one side of the coal pillar goaf 2, the free surface of the initial borehole 4 is only the roof fracture surface on one side of the goaf 2, which severely restricts the blasting effect. To optimize the arrangement of boreholes 4, the density of blast holes is increased within the first 50m of the borehole arrangement. Preferably, the spacing between boreholes 4 is less than 50cm, so as to form significant large cracks along the direction of the borehole arrangement, promote the rapid collapse of the roof, and form a new blasting free surface, which is conducive to the blasting effect of subsequent blast holes. The spacing of subsequent blast holes can be gradually increased according to the blasting situation. During the blasting process, in order to increase the directionality of the detonation wave propagation, blank reserved holes 14 are set before and after the blasting boreholes 4 for blasting.

[0059] To achieve roof cutting on one side of the thick, hard roof in the coal pillar goaf 2 during top coal caving mining and eliminate the influence of the triangular zone, borehole 4 is first drilled in the coal pillar roadway 1. The inclination angle θ of borehole 4 is determined by combining the thickness a of the coal seam roof 3 and the width b of the coal pillar, with tanθ = a / b. The preferred borehole angle is 45-55°. After borehole 4 is completed, a single-launcher, dual-receiver sonic logging instrument is used. Based on the different elasticity, density, and fluid properties in different rocks, the sound wave propagation speed varies, and the received sound wave time difference is used to determine the rock strata properties and distribution. Then, based on the lithology of the rock strata, the spacing between adjacent boreholes 4, and the properties of the explosive, the charge amount per unit length of the borehole is determined, and the explosive cartridge diameter and density are determined based on the charge amount per unit length. Of course, the lithology and distribution of the drilled rock strata can also be obtained using other existing technologies, all of which are well-known and will not be elaborated here.

[0060] Unless otherwise stated, if any of the technical solutions disclosed in this invention specify a numerical range, then the disclosed numerical range is a preferred numerical range. Anyone skilled in the art should understand that the preferred numerical range is merely one among many feasible numerical values ​​that has a more obvious or representative technical effect. Because there are many numerical values, it is impossible to list them all. Therefore, this invention discloses only some numerical values ​​to illustrate the technical solutions of this invention. Furthermore, the numerical values ​​listed above should not constitute a limitation on the scope of protection of this invention.

[0061] Furthermore, if the present invention discloses or relates to mutually fixedly connected components or structural parts, then unless otherwise stated, a fixed connection can be understood as: a detachable fixed connection (e.g., using bolts or screws), or a non-detachable fixed connection (e.g., riveting, welding). Of course, mutually fixed connections can also be replaced by an integral structure (e.g., manufactured using a casting process) (except where it is obviously impossible to use an integral molding process).

[0062] Furthermore, unless otherwise stated, the terms used to indicate positional relationships or shapes in any of the technical solutions disclosed in this invention include states or shapes that are similar to, analogous to, or close to those states or shapes. Any component provided by this invention can be assembled from multiple individual components or can be a single component manufactured using a one-piece molding process.

[0063] The above embodiments are merely illustrative examples to clearly illustrate the present invention and are not intended to limit the implementation. Those skilled in the art will recognize that other variations or modifications can be made based on the above description. It is neither necessary nor possible to exhaustively list all embodiments here. However, obvious variations or modifications derived therefrom are still within the scope of protection of this invention.

Claims

1. A method for blasting and cutting the roof to relieve pressure in the overlying strata of the goaf on the side of the coal pillar in a top-coal caving working face, characterized in that: In the coal pillar roadway (1), an inclined borehole (4) is drilled on the coal seam roof (3) towards the goaf (2), and the borehole (4) penetrates the hard rock layer (5) of the coal seam roof (3). The borehole (4) is arranged in front of the advanced stress concentration area of ​​the coal seam roof (3) and distributed from dense to sparse along the advancing direction of the coal mining face (6). Then, decoupled charges are carried out in each borehole (4), and detonation is started from the dense area of ​​borehole (4) to carry out advanced pre-splitting of the coal seam roof (3) above the goaf (2). Afterwards, as the coal mining face (6) advances, the advanced stress concentration area of ​​the coal seam roof (3) moves forward continuously, thereby causing the coal seam roof (3) above the goaf (2) to collapse further, so as to realize the blasting and roof cutting pressure relief of the overlying rock layer of the goaf (2). The explosive charge structure inside the borehole (4) includes a shaped charge (8) placed in the hard rock layer (5) section and a regular charge (9) placed in the soft rock layer (11) section. The outermost end of the borehole (4) is sealed by a plug (16). The shaped charge (8) includes a flexible metal mesh cylinder (801), particulate explosive (802) filled in the flexible metal mesh cylinder (801), and a spreading assembly that causes the flexible metal mesh cylinder (801) to axially contract and radially expand so that it abuts against the wall of the borehole (4). The spreading assembly includes a telescopic rod (803) disposed inside the flexible metal mesh cylinder (801). The two ends of the telescopic rod (803) are fixedly connected to the two ends of the flexible metal mesh cylinder (801). The telescopic rod (803) is also provided with a support frame (804) for spreading the two ends of the flexible metal mesh cylinder (801). The telescopic rod (803) includes an upper tube section (8031) and a lower tube section (8032). The upper end of the lower tube section (8032) is slidably fitted outside the upper tube section (8031). One end of the support frame (804) is hinged to the lower tube section (8032), and the other end is hinged to the flexible metal mesh cylinder (801). The upper tube section (8031) is provided with a spring pin (805), and the inner wall of the lower tube section (8032) is provided with a pin hole (806) that cooperates with the spring pin (805). During loading, the shaped charge (8) is fed into the borehole (4) using a push rod until it comes into contact with the ordinary charge (9) loaded in the soft rock layer (11) above. Then the push rod continues to push the shaped charge (8), causing the lower pipe section (8032) to move upward relative to the upper pipe section (8031). This forces the support frame (804) to open the flexible metal mesh cylinder (801) while continuously squeezing and pushing the granular explosive (802) inside the flexible metal mesh cylinder (801), so that the flexible metal mesh cylinder (801) abuts against the wall of the borehole (4) until the spring pin (805) is inserted into the pin hole (806). The telescopic rod (803) and the flexible metal mesh cylinder (801) are also connected by several pull ropes (807) that allow the side wall portion of the flexible metal mesh cylinder (801) to detach from the wall of the borehole (4); The telescopic rod (803) is a tube made of iron filings, and the tube is filled with a single-element high explosive (13). When detonated, the single-element high explosive (13) detonates the surrounding particulate explosives (802) at the same time. One support frame (804) is provided at the upper end and one at the lower end of the lower pipe section (8032). Each support frame (804) includes multiple connecting rods (8041) evenly arranged along the circumference of the lower pipe section (8032). One end of each connecting rod (8041) is hinged to the lower pipe section (8032), and the other end extends forward and outward to be hinged to the flexible metal mesh cylinder (801). The lower end face of the flexible metal mesh cylinder (801) is connected to the connecting rods (8041) of the support frame (804) located below by multiple cable ties.

2. The method for blasting and cutting the roof to relieve pressure in the overlying strata of the coal pillar side goaf in a top-coal caving working face according to claim 1, characterized in that: The borehole (4) is equipped with staged energy reduction end caps (10) at both ends of the charge structure. One staged energy reduction end cap (10) is located at the top of the borehole (4), and the other staged energy reduction end cap (10) is located between the charge structure and the plug (16) and above the coal-rock interface.

3. The method for blasting and cutting the roof to relieve pressure in the overlying strata of the coal pillar side goaf in a top-coal caving working face according to claim 2, characterized in that: The graded energy reduction end (10) includes two natural rubber substrates (101) arranged vertically and connected by a connecting rod, and multiple metal material layers (102) spaced apart on the connecting rod. Each metal material layer (102) has an outwardly convex conical surface facing the outer surface of the charge structure.

4. The method for blasting and cutting the roof to relieve pressure in the overlying strata of the coal pillar side goaf in a top-coal caving working face according to any one of claims 1-3, characterized in that: A blank reserved hole (14) without charging is provided between any two adjacent boreholes (4).

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