CN112179228A - Deep hole subsection blasting joint cutting control top plate overall collapse method - Google Patents
Deep hole subsection blasting joint cutting control top plate overall collapse method Download PDFInfo
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- F—MECHANICAL ENGINEERING; LIGHTING; HEATING; WEAPONS; BLASTING
- F42—AMMUNITION; BLASTING
- F42D—BLASTING
- F42D1/00—Blasting methods or apparatus, e.g. loading or tamping
- F42D1/08—Tamping methods; Methods for loading boreholes with explosives; Apparatus therefor
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- F42—AMMUNITION; BLASTING
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Abstract
A method for controlling the whole collapse of a top plate by a deep hole subsection blasting kerf belongs to the technical field of blasting kerf and can solve the problem that the top plate of a coal seam hangs but does not fall, so that the mine pressure is strongly developed, the deep hole blasting is carried out on an old top rock stratum to prevent the top plate from collapsing immediately after blasting, the old top rock stratum in a blasting section after blasting can form a pre-crack weak surface with the same inclination angle as the natural collapse angle of the rock stratum along a drilling arrangement line, the structural integrity of the old top rock stratum is cut off, the top rock stratum is directionally collapsed along the pre-crack weak surface under the action of gravity, the concentrated stress on a stoping roadway and the front part of a coal wall caused by the large-area overhang of the top plate can be effectively reduced, the stress environment of the roadway is improved, and the top plate behind a working face collapses in time, so that the air leakage, air leakage and air leakage of a goaf are avoided, The gas is gathered and the spontaneous combustion of the residual coal occurs.
Description
Technical Field
The invention belongs to the technical field of blasting joint cutting, and particularly relates to a treatment method for controlling the whole caving of a top plate by deep-hole segmented blasting joint cutting, wherein the deep-hole segmented blasting joint cutting causes strong mine pressure appearance of a working face due to the fact that a coal mine roadway-side top plate cantilever structure, a working face end top plate, a goaf top plate in a primary mining stage and a final mining frame-removing channel top plate are suspended without falling.
Background
The phenomenon that a roof is not prone to collapse along with mining exists in the coal mining process of a coal mine, large-area overhanging of the roof behind a goaf of a working face can be caused along with the pushing of the working face, so that concentrated stress borne by the front of a stoping roadway and a coal wall is increased, air leakage of the goaf is serious, and potential safety hazards of gas gathering and spontaneous combustion and ignition of residual coal in the goaf are generated. In order to solve the problem, passive methods such as reinforcing support, reserving wider roadway protection coal pillars and the like are mostly adopted to maintain the stability of the stoping roadway, the method cannot solve the problem from the source, the engineering quantity is large, and the waste of coal resources is caused. Based on the method, the method for controlling the whole collapse of the top plate by arranging the deep hole blasting drilling and the sectional blasting joint cutting on the coal seam top plate is provided.
In the published literature, a patent of 'a coal face end mining withdrawal channel top cutting pressure relief roadway protection method' with publication number CN 108661643 a can be obtained, and the method for coal seam roof top cutting pressure relief is to uniformly arrange deep hole blasting drill holes along a support of a digging working face and a top plate of a withdrawal channel of other withdrawal equipment, and from a position 10m away from one end of the withdrawal channel from the end, a deep hole pre-splitting drill hole is arranged at the middle position of the top plate in the withdrawal channel, all the drill holes are deviated to 45 degrees from the beginning drilling end of the withdrawal channel, namely, the included angle between the drill holes and the vertical direction is 45 degrees, and a propagation path of the mining advance supporting pressure of the working face is cut off through the deep hole pre-splitting coal seam roof. The method has at least two defects or shortcomings, one is that deep hole blasting drill holes are uniformly distributed along a support of the excavation working face and a top plate of a withdrawal channel of other mining equipment, the number of the drilled deep hole blasting drill holes is large, the engineering quantity is large, the explosive consumption is large when explosive blasting is subsequently carried out, and the danger degree is increased; secondly, the included angle between the hole and the vertical direction is 45 degrees, the determination of the drilling elevation angle is not scientific, the top plate collapses along 45 degrees after blasting and roof cutting, and the pressure relief effect is poor.
In the patent of 'a hard roof deep hole presplitting blasting roof cutting pressure relief method' with publication number of CN 103278055A, the arrangement method of roof cutting pressure relief blasting drill holes is that blasting holes and control holes are uniformly arranged at intervals in the same plane above the roof of a special roof cutting roadway or a stoping roadway, the distance between the adjacent blasting holes and the control holes is 5m, and roof cutting blasting is carried out after charging and roof cutting blasting in the blasting holes to break the hard roof along cracks generated by presplitting blasting, so that the aim of roof cutting pressure relief is achieved. The method has at least two defects or shortcomings, one is that when the blast holes and the control holes are uniformly arranged at intervals, stress disturbance generated when drilling the holes is easy to collapse adjacent drilled holes, so that the blasting holes are difficult to charge, the charge amount is reduced, and the pressure relief effect is reduced; and secondly, blasting drill holes in the same plane are positioned right above the roadway, and sudden collapse of a roof rock stratum during blasting roof cutting may cause sudden increase of the overburden pressure borne by the roadway roof, so that potential safety hazards exist.
In order to solve the problems that the concentrated stress on the front of a mining roadway and a coal wall is increased and the mine pressure is strongly shown due to the fact that a hard top plate of the roadway is suspended in a large area and does not fall, and the purpose that the stress environment of surrounding rocks is improved due to the fact that the top plate rock stratum is directionally collapsed on the premise that the safety and the stability of a blasting operation roadway are guaranteed, a method for controlling the whole collapse of a top plate by a deep-hole subsection blasting joint of the hard top plate of the mining roadway is urgently needed.
Disclosure of Invention
The invention aims to solve the problem that a coal seam roof is suspended but not fallen, ensure that the coal seam roof can be collapsed along with mining, and further reduce the stress concentration phenomenon of a coal wall and a roadway of a working face caused by overlarge suspended roof area, thereby effectively controlling the deformation of roadway surrounding rocks, achieving the effect of ensuring the stability of the roadway surrounding rocks, and providing the method for controlling the integral collapse of the roof by cutting seams in the deep-hole sectional blasting.
The invention adopts the following technical scheme:
a deep hole subsection blasting joint cutting control top plate overall caving method comprises the following steps:
firstly, determining the height H of blasting top cutting according to the distance between the top of an old top rock stratum and a roadway top plate;
the second step, in the deep hole blasting, the deep hole blasting is divided into a charging section and a hole sealing section, wherein the vertical height of the charging section is the thickness H of the old roof rock layer in the coal seam roof1Vertical height H of hole sealing section in deep hole blasting2Is the difference between the height of the top cut and the vertical height of the charge segment, i.e. H2=H-H1；
Thirdly, determining the length L of the whole collapse of the top plate needing deep-hole sectional blasting joint cutting control, wherein the L is the length of the collapse of the top plate under blasting control;
fourthly, when the uniaxial compressive strength of the old roof rock stratum is more than 80MPa, the interval L of the adjacent deep hole blasting drill holes is formed when the sectional blasting joint cutting controls the whole collapse of the roof1Is 6 m; when the uniaxial compressive strength of the old roof rock stratum is 60-80 MPa, the distance L between adjacent deep hole blasting drill holes is controlled by the sectional blasting joint cutting to control the whole collapse of the roof1Is 5 m; when the uniaxial compressive strength of the old roof rock stratum is less than 60MPa, the distance L between the adjacent deep hole blasting drill holes is controlled by the sectional blasting joint cutting when the whole roof collapses1Is 4 m;
fifthly, designing an L2-3 times of one end of the area to be blasted1Arranging a first deep hole blasting drill hole at a distance of 2-3 times L from the other end head1Arranging the last deep hole blasting drill hole;
sixthly, taking every 5 deep hole blasting drill holes as one section, wherein the 3 rd deep hole blasting drill hole in the middle of each section is a spare drill hole and is used for guiding the direction of blasting joint cutting, if the rest 4 deep hole blasting drill holes in the section have hole collapse phenomenon before charging, the charging is difficult, the blasting drill holes with the hole collapse phenomenon are abandoned, and the charging is carried out to the 3 rd spare deep hole blasting drill hole in the middle of the section;
seventhly, calculating the section spacing L of two adjacent sections of deep hole blasting drill holes2，
Wherein H is1Is the thickness of the old roof rock layer, RtIs the tensile strength of the old roof rock stratum, sigmacRadial stress generated by the center of the explosive charging section of the blast hole;
eighthly, arranging a plurality of deep hole blasting drill holes in the direction of a top plate of a to-be-blasted area to the rock stratum caving space at an elevation angle of 72 degrees under the condition of blasting operation, wherein the depth of each deep hole blasting drill hole is H = H/sin72 degrees, and the diameter of each deep hole blasting drill hole is phi =94mm, and numbering the drill holes according to the sequence of the drilling;
ninthly, after the deep hole blasting drilling is finished, watering and dust falling are carried out on the roadway in the range of 80m in front of and behind the operation point, then 4 blasting drilling holes in each section are filled with explosives, and the drilling holes are checked before loading the explosives to ensure that the explosive is wound on the hole bottom;
step ten, calculating the length of the old top rock stratum charging section aimed at in the deep hole blasting drilling to be h1=H1/sin72 DEG, the length of the hole sealing section is h2=H2/sin72°；
Step eleven, filling explosives into the deep hole blasting holes in a non-coupling explosive filling mode, wherein the non-coupling coefficient of radial explosive filling is 94/63= 1.49; when a first explosive roll is loaded, two detonating cords are bent inwards to the explosive through the end head of the first explosive roll and inserted into the explosive roll by 25cm, the detonating cords at the end head of the first explosive roll are fixed by using an adhesive tape and then sent to the bottom of a hole, the detonating cords are prevented from falling off in the pushing process of the explosive roll, then other explosive rolls are sequentially loaded into the blasting drill hole, the explosive loading length reaches the explosive loading length h1Then carrying out hole sealing operation, and filling the hole sealing stemming in the same wayFilling the hole sealing section;
step ten, two detonating cords in each deep hole blasting drill hole are detonated by coal mine allowed electric detonators, namely double detonators and double detonating cords are adopted for detonating, 2 detonators are connected in parallel outside the hole, and blasting cords have to be well insulated and hung in the air to avoid touching the ground;
step thirteen, serially connecting blasting lines suspended in 4 deep hole blasting drill holes in each section, connecting a blast line in each section in a mode of single-hole parallel connection and four-hole serial connection, and spraying water and reducing dust for the roadway within the range of 80m of an operation point after the conditions of support, gas, ventilation and the like of the work place are checked and the safety is ensured, so that blasting can be realized;
fourteenth, using a coal mine allowable blasting device, adopting forward detonation, detonating blast holes (4) in one blasting section each time, sequentially detonating N deep hole blasting sections in N times, rechecking the conditions of support, gas, ventilation and the like after blast smoke generated by the previous blasting is blown away to ensure safety, flushing coal dust in a range of 80m in front of and behind the operation site, and then carrying out next blasting;
fifthly, after blasting of all the charging blast holes is finished, when blast smoke is blown away completely, and after the blast smoke is exhausted (at least 15 min), the blasting worker, the team leader and the tile inspector firstly enter the lower part of the bracket to inspect the conditions of ventilation, gas, support, top plate and the like, hang a metal net in the blasted area and strictly forbid the entry of the personnel;
sixthly, after the deep hole sectional blasting operation is finished for at least one shift (8 hours), the production operation can be carried out.
The deep hole blasting drill holes are located on the same straight line.
The 3 rd blasting drill hole in the middle of each section of deep hole blasting drill hole is a standby drill hole, generally does not charge, is used for guiding the direction of blasting joint cutting, and can be used as a blasting drill hole for carrying out charging blasting when the rest 4 blasting drill holes have hole collapse during charging.
The elevation angle of the deep hole blasting drilling hole is 72 degrees determined by a natural collapse angle of a hard rock stratum, and the top cutting effect can be improved by combining a pre-crack weak surface generated by blasting.
The explosive used for blasting is a three-level coal mine allowable emulsion explosive, the specification of the explosive cartridge is phi 63 multiplied by 800mm, and the density is 1194.27kg/m32.0 Kg/roll, the detonator is a No. 9 common instantaneous electric detonator allowed for coal mines, the detonating cord is a detonating cord allowed for coal mines, the specification is phi 5.2-5.5mm, and the detonation speed is 2800 m/s.
The number of the emulsion explosive cartridges filled in the explosive charging length is an integer, and when the residual explosive charging length is smaller than 800mm of the cartridges in the explosive charging process, a cartridge of emulsion explosive is filled at last and then hole sealing is carried out, so that the blasting effect is ensured.
And each deep hole blasting drill hole filled with explosives is provided with double detonators and double detonating cords, so that blasting work is guaranteed to be absolutely lost.
The hole sealing stemming is controlled to be 8-15% in humidity and 5-15 cm in length, so that a good hole sealing effect is guaranteed.
The detonator is a BF-200 type detonator, charges once, explodes in sections in sequence, detonates 4 blasting drill holes in the first section at the same time, after the crack in the rock body is expanded stably under the action of detonation wave during the first detonation, detonates 4 blasting drill holes in the second section at the same time after blasting smoke is blown away, ventilation and gas are recovered to be normal, and finishes blasting operation of all sections in sequence, namely, detonates N deep hole blasting sections in sequence for N times.
Blasting the old roof rock stratum; and arranging deep hole blasting drill holes in N sections, reserving a spare drill hole for guiding the direction of blasting cutting joints in each section, and sequentially detonating the deep hole blasting drill holes in each section for N times to make the roof rock layer collapse along the pre-crack weak surface under the action of gravity.
The invention has the following beneficial effects:
the technical scheme provided by the invention solves the problem that the roof of the coal mining coal seam hangs but does not fall to cause the strong mine pressure, can prevent the roof from caving immediately after blasting by carrying out deep hole blasting on the old roof rock stratum, the old roof rock stratum in the blasting section after blasting can form a pre-crack weak face with the same inclination angle as the natural collapse angle of the rock stratum along a drilling arrangement line, and the integrity of the structure of the old roof rock stratum is cut off, so that the roof rock stratum directionally collapses along the pre-crack weak face under the action of gravity, the concentrated stress on the front of a stoping roadway and a coal wall caused by large-area roof suspension of the roof can be effectively reduced, the stress environment of the roadway is improved, meanwhile, the roof behind a working face collapses in time, and the occurrence of the conditions of air leakage, gas accumulation and spontaneous combustion of residual coal is avoided.
Compared with the prior art, the method carries out sectional deep hole blasting on the old roof rock stratum in the coal seam roof, the elevation angle of the blasting drilling hole is determined to be 72 degrees in combination with the natural collapse angle of the rock stratum, the direct roof rock stratum is not processed, the arrangement quantity of the blasting drilling holes and the using amount of the mining emulsion explosive are reduced, the collapse difficulty of the roof rock stratum is reduced, and the effect of overall collapse of the roof is achieved. After the first section of blasting is finished, main cracks generated by the blasting can be expanded along the direction of a drilling arrangement line, cracks generated during the second section of blasting can be communicated with cracks generated by the first section of blasting, after the sectional sequential blasting is finished, the old top rock stratum in the blasting section forms a pre-crack weak face with the same inclination angle as the natural collapse angle of the rock stratum along the drilling arrangement line, the integrity of the structure of the old top rock stratum is cut off, the directional collapse of the top plate is controlled under the action of the gravity of the top plate rock stratum, the stress environment of the roadway can be effectively improved, and the conditions of air leakage, gas gathering and spontaneous combustion of residual coal in a goaf are avoided.
Drawings
FIG. 1 is a front view of the arrangement of the boreholes in a face-side open-off roadway of the present invention.
FIG. 2 is a plan view of the arrangement of the boreholes in the face-side open-off roadway of the present invention.
FIG. 3 is a left side view of the hole pattern in the face cut roadway of the present invention.
FIG. 4 is a schematic diagram of explosive cartridge, detonating cord, detonator and sealing in the blasting borehole according to the present invention.
Fig. 5 is a partially enlarged view of a portion a of fig. 4.
Fig. 6 is a partially enlarged view of fig. 4 at B.
Fig. 7 is a sectional blasting on-line diagram of the present invention.
FIG. 8 is a schematic of crack propagation after a single deep hole blast drilling blast.
Fig. 9 is a diagram of the slitting effect of deep hole sectional blasting.
Wherein: 1-old top of coal bed; 2-directly jacking the coal bed; 3-coal bed; 4-cutting solid coal behind the holes; 5-transporting the crossheading; 6-return air crossheading; 7-stope face; 8-borehole elevation; 9-deep hole blasting and drilling; 10-drilling for standby; 11-length of the medicine section in the drill hole; 12-length of the hole sealing section in the drilled hole; 13-deep hole blasting and drilling hole spacing; 14-deep hole blasting drilling section spacing; 15-controlling the caving length of the top plate by blasting; 16-allowable emulsion explosive for coal mine; 17-stemming; 18-allowable detonating cords for coal mines; 19-a first coil of explosive inner detonating cord; 20-a first coil of explosive external detonating cord fixing strip; 21-coal mine allowable electric detonator; 22-electric detonator leg wire; 23-blasting line; 24-coal mine allowed exploder; 25-an expansion area; 26-a crushing zone; 27-a fracture zone; 28-main crack.
Detailed Description
The present invention will be described in further detail with reference to the accompanying drawings.
Examples
The average thickness of a coal seam 3 mined from a certain mine is 2.45m, the average dip angle is 4 degrees, an old top 1 of the coal seam is fine-grained sandstone with the average thickness of 6m, a direct top 2 of the coal seam is mudstone with the average thickness of 6.65m, the initial caving step distance of a top plate during mining of an upper section working face is about 35m, the period pressure step distance is about 20m, the ore pressure is strongly shown, and the deformation of a leading influence section of a mining roadway is serious, so that a deep hole subsection blasting kerf control top plate integral caving method is used in a lower section working face open-cut roadway to reduce the initial caving step distance of the working face.
The distance between the old top of the coal seam and the top plate of the roadway is 12.65m, and the blasting roof cutting height H =12.65m is determined; the average thickness of the old roof rock layer is 6m, so that the vertical height H of the charging section in the deep hole blasting drilling is determined1=6m, vertical height H of the hole sealing section2=12.65-6=6.65m。
The working face inclined length is 164m, and the blasting control roof collapse length 15 is 164 m.
And determining that the hole spacing 13 of deep hole blasting drilling is 5m because the uniaxial compressive strength of the old-top fine-grained sandstone of the working surface is 78 MPa.
The tensile strength of the working face old top fine sandstone is 3.1MPa, and the axial coefficient of charge is lc1, pressure increase coefficient n of 10, old jackThe poisson ratio mu of the rock stratum is 0.3, and the formula is substituted
And obtaining the distance between the deep hole blasting drill sections of 9.5m and the number N = (L-2 multiplied by 2.5L) of the designed deep hole blasting sections1）/（4L1+L2) And = 164-25)/(20 + 9.5) =4.7, so that five sections of blast drill holes can be arranged in the working face cutting hole, the first blast drill hole is arranged at a position 13m from the transportation crossheading 5, the last blast drill hole is arranged at a position 13m away from the return air crossheading 6, five sections of blast drill holes are arranged, and each 5 drill holes are one section.
The deep hole blasting drill holes 9 are arranged 13m from the transportation gate way 5, the deep hole blasting drill holes 9 are arranged in the direction of a drilling elevation angle 8 of 72 degrees towards the rock stratum caving space of a top plate of a region to be blasted, the drilling depth is H = H/sin72 degrees =12.65/sin72 degrees =13.3m, the drilling hole diameter is phi =94mm, and the drilling holes are numbered according to the drilling sequence, wherein the numbers are 1#, 2#, 3#, 4#, 5# 6321 #, 22#, #23, #24 and # 25 #.
The 1#, 2#, 3#, 4#, and 5# deep hole blasting drill holes 9 are first sections, the 6#, 7#, 8#, 9#, and 10# deep hole blasting drill holes 9 are second sections, the 11#, 12#, 13#, 14#, and 15# deep hole blasting drill holes 9 are third sections, the 16#, 17#, 18#, 19# and 20# deep hole blasting drill holes 9 are fourth sections, the 21#, 22#, 23#, 24# and 25# deep hole blasting drill holes 9 are fifth sections, the 3#, 8#, 13#, 18# and 23# are respectively spare drill holes 10 in the first sections, the second sections, the third sections, the fourth sections, and the fifth sections, and no powder is filled in the spare drill holes 10, and the rest of the deep hole blasting drill holes 9 are filled with powder.
After the deep hole blasting drilling is completed, the roadway within the range of 80m in front of and behind the operation point is sprayed with water for dust reduction, then the rest 4 blasting drilling holes except the spare drilling hole 10 in each section are prepared for charging, the drilling holes are checked before charging, the explosive package is ensured to be loaded to the hole bottom, and the spare drilling holes are not charged.
Calculating the length 11 of the explosive section in the old roof rock stratum drilling hole aiming at the deep hole blasting drilling hole to be h1=H1The angle of/sin 72 degrees is not less than 6.3m, the length of the hole sealing section is 12 h2=H2/sin72°=7m。
Loading three-level coal mine allowable emulsion explosive 16 with the specification of phi 63 multiplied by 800mm into a deep hole blasting drill hole 9 by adopting a non-coupling explosive loading mode, when a first coil of explosive is loaded, bending two coal mine allowable explosive cables 18 towards the explosive through the end head of the first coal mine allowable emulsion explosive 16 to ensure that the length of the explosive inserted into the first coil of explosive internal explosive cable 19 is 25cm, winding the first coil of explosive internal explosive cable 19 outside the first coal mine allowable emulsion explosive 16 by using adhesive tapes to prepare a first coil of explosive external explosive cable fixing band 20 for preventing the falling of the first coil of explosive internal explosive cable 19 in the explosive coil pushing process, and then sequentially loading other coal mine allowable emulsion explosive 16 into the blasting drill hole 9,
the length of the explosive section 11 in the drill hole is 6.3m, the length of the explosive roll of the allowable emulsion explosive 16 for the three-level coal mine is 800mm, therefore, after 8 rolls of emulsion explosive are filled in each deep hole blasting drill hole 9, hole sealing operation is carried out, and the hole sealing stemming is filled in the hole sealing length 12 in the same way.
Two detonating cords in each deep hole blasting drill hole 9 are detonated by coal mine allowable electric detonators 21, namely double detonators and double detonating cords are adopted for detonation, 2 detonators are connected in parallel outside the hole, and a blasting line 23 must have good insulation and be hung in the air to prevent the blasting line from touching the ground.
And (3) loading the 20 deep hole blasting drill holes 9 and carrying out online operation by using the blasting line 23 after hole sealing is finished, connecting two electric detonators in the hole openings of the deep hole blasting drill holes 9 after the loading is finished in each section in parallel, and connecting 4 deep hole blasting drill holes 9 after the loading is finished in each section in series so as to ensure that the 4 deep hole blasting drill holes 9 in the same section are simultaneously detonated. After the five sections of deep hole blasting sections are connected on line, the conditions of supporting, gas, ventilation and the like of a working site are checked to ensure safety, and then the roadway within the range of 80m of the working site is sprayed with water again to reduce dust so as to be blasted.
The method is characterized in that a BF-200 type detonator allowed for coal mines is used, positive detonation is adopted, blast holes (4) in one blasting section are detonated each time, five deep hole blasting sections are detonated in sequence 5 times, after blast smoke generated by the last blasting is blown away, the conditions of support, gas, ventilation and the like are checked again to ensure safety, coal dust in the range of 80m around an operation site is washed, and next blasting can be carried out.
After blasting of the total five sections of deep hole blasting sections is completed, waiting for blasting smoke to be blown away completely, checking seriously to confirm that no misfiring blasting exists, after 15min, blasting workers, team leaders and tile inspectors firstly enter the lower part of the support to check the conditions of ventilation, gas, support, top plate and the like, and hang a metal mesh in a blasted area to strictly forbid personnel to enter.
After the blasting operation is finished for 8 hours, the working face carries out stoping operation, the working resistance of the hydraulic support is displayed within 30 days, after the deep-hole segmented blasting joint cutting is used for controlling the whole collapse of the top plate to carry out old-roof blasting pre-splitting, the initial collapse step of the working face is reduced to 20m, the periodic pressure step is reduced to 13m, the initial collapse step and the periodic pressure step are respectively reduced by 43% and 35% compared with those of other working faces, surrounding rocks of the stoping roadway in the advance influence section are not obviously deformed, the goaf is basically filled after the top plate at the rear of the goaf is observed, no gas overrun phenomenon occurs, and the deep-hole segmented blasting joint cutting control method for the whole collapse of the top plate achieves good effect.
Claims (3)
1. The utility model provides a deep hole segmentation blasting joint-cutting control roof whole caving method which characterized in that: the method comprises the following steps:
the method comprises the steps that firstly, the position of an old top in a top plate of a coal seam to be mined is determined, the old top is a rock stratum which is thick above the coal seam, hard and not easy to collapse along with mining, when the old top is suspended in a large area, the pressure of a working face is increased rapidly, a hydraulic support is pressed, and the blasting roof cutting height H is the distance between the top of a rock stratum with the old top and a roadway top plate;
the second step, in the deep hole blasting, the deep hole blasting is divided into a charging section and a hole sealing section, wherein the vertical height of the charging section is the thickness H of the old roof rock layer in the coal seam roof1Vertical height H of hole sealing section in deep hole blasting2Is the difference between the height of the top cut and the vertical height of the charge segment, i.e. H2=H-H1；
Thirdly, determining the length L of the whole collapse of the top plate needing deep-hole sectional blasting joint cutting control, wherein the L is the length of the collapse of the top plate under blasting control;
fourthly, determining the distance L between adjacent deep hole blasting drill holes when the sectional blasting joint cutting control top plate integrally collapses according to the uniaxial compressive strength of the old top rock stratum1；
Fifthly, designing an L2-3 times of one end of the area to be blasted1Arranging a first deep hole blasting drill hole at a distance of 2-3 times L from the other end head1Arranging the last deep hole blasting drill hole;
sixthly, dividing each 5 deep hole blasting drill holes into one section, wherein the 3 rd deep hole blasting drill hole in the middle of each section is a standby drill hole and is used for guiding the direction of blasting cutting joint;
seventhly, calculating the section spacing L of two adjacent sections of deep hole blasting drill holes2，
Wherein H is1Is the thickness of the old roof rock layer, RtIs the tensile strength of the old roof rock stratum, sigmacRadial stress generated by the center of the explosive charging section of the blast hole;
eighthly, arranging a plurality of deep hole blasting drill holes in the direction of a top plate of a to-be-blasted area to the rock stratum caving space at an elevation angle of 72 degrees under the condition of blasting operation, wherein the depth of each deep hole blasting drill hole is H = H/sin72 degrees, and the diameter of each deep hole blasting drill hole is phi =94mm, and numbering the drill holes according to the sequence of the drilling;
ninthly, after the deep hole blasting drilling is finished, watering and dust falling are carried out on the roadway in the range of 80m in front of and behind the operation point, and then 4 blasting drilling holes in each section are filled with explosives;
step ten, calculating the length of the old top rock stratum charging section aimed at in the deep hole blasting drilling to be h1=H1/sin72 DEG, the length of the hole sealing section is h2=H2/sin72°；
Tenth itemFilling explosives into the deep-hole blasting hole in a non-coupling explosive filling mode, wherein the non-coupling coefficient of radial explosive filling is 94/63= 1.49; when a first explosive roll is loaded, two detonating cords are bent inwards to the explosive through the end head of the first explosive roll, inserted into the explosive roll by 25cm, fixed by adhesive tape and then sent to the bottom of a hole, and then other explosive rolls are loaded into the blasting drill hole in sequence, wherein the explosive loading length reaches the explosive loading length h1Then hole sealing operation is carried out, and hole sealing stemming is filled in a hole sealing section in the same way;
step ten, two detonating cords in each deep hole blasting drill hole are detonated by coal mine allowable electric detonators, namely double detonators and double detonating cords are adopted for detonation, 2 detonators are connected in parallel outside the hole, and blasting wires are insulated and hung in the air;
step thirteen, serially connecting blasting lines suspended in 4 deep hole blasting drill holes in each section, connecting a blast line in each section in a mode of single-hole parallel connection and four-hole serial connection, and spraying water and reducing dust for a roadway within 80m of an operation point after safety of support, gas and ventilation conditions of a work place are checked, so that blasting can be performed;
fourthly, using a permitted blasting device for the coal mine, adopting positive detonation, detonating a blast hole in one blasting section each time, sequentially detonating N deep-hole blasting sections in N times, checking the support, gas and ventilation conditions again after blasting smoke generated by the previous blasting is blown away to ensure safety, and flushing coal dust in a range of 80m in front of and behind the operation site so as to carry out next blasting;
fifthly, after blasting of all the charging blast holes is finished, when blasting smoke is completely blown away, checking to confirm that no misfiring explosion exists, and after the blasting smoke is completely exhausted, entering the lower part of the bracket, checking the conditions of ventilation, gas, support and top plate, and hanging a metal net in a blasted area;
sixthly, after the deep hole sectional blasting operation is finished for at least one shift, the production operation can be carried out.
2. The deep hole subsection blasting joint-cutting control top plate overall caving method according to claim 1, characterized in that: fourth step the said LaoWhen the uniaxial compressive strength of the top rock stratum is more than 80MPa, the distance L between the adjacent deep hole blasting drill holes is controlled by the sectional blasting cutting joint when the top plate is wholly collapsed1Is 6 m; when the uniaxial compressive strength of the old roof rock stratum is 60-80 MPa, the distance L between adjacent deep hole blasting drill holes is controlled by the sectional blasting joint cutting to control the whole collapse of the roof1Is 5 m; when the uniaxial compressive strength of the old roof rock stratum is less than 60MPa, the distance L between the adjacent deep hole blasting drill holes is controlled by the sectional blasting joint cutting when the whole roof collapses1Is 4 m.
3. The deep hole subsection blasting joint-cutting control top plate overall caving method according to claim 1, characterized in that: the deep hole blasting drill holes are located on the same straight line.
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