CN112647945A - Hard roof cutting method for protective layer mining - Google Patents
Hard roof cutting method for protective layer mining Download PDFInfo
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- CN112647945A CN112647945A CN202011538593.9A CN202011538593A CN112647945A CN 112647945 A CN112647945 A CN 112647945A CN 202011538593 A CN202011538593 A CN 202011538593A CN 112647945 A CN112647945 A CN 112647945A
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- 238000005520 cutting process Methods 0.000 title claims abstract description 69
- 238000005065 mining Methods 0.000 title claims abstract description 27
- 238000000034 method Methods 0.000 title claims abstract description 26
- 239000011241 protective layer Substances 0.000 title claims abstract description 16
- 238000005422 blasting Methods 0.000 claims abstract description 28
- 239000003245 coal Substances 0.000 claims abstract description 22
- 239000011435 rock Substances 0.000 claims abstract description 13
- 230000000694 effects Effects 0.000 claims abstract description 11
- 239000010410 layer Substances 0.000 claims abstract description 6
- 238000005553 drilling Methods 0.000 claims description 27
- 238000007789 sealing Methods 0.000 claims description 17
- 239000003814 drug Substances 0.000 claims description 11
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Substances O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 claims description 11
- 238000010276 construction Methods 0.000 claims description 10
- 238000005086 pumping Methods 0.000 claims description 7
- 239000002893 slag Substances 0.000 claims description 6
- 238000007689 inspection Methods 0.000 claims description 5
- 230000008859 change Effects 0.000 claims description 4
- 239000004568 cement Substances 0.000 claims description 3
- 238000012544 monitoring process Methods 0.000 claims description 3
- 239000003566 sealing material Substances 0.000 claims description 3
- 239000002002 slurry Substances 0.000 claims description 2
- 238000005474 detonation Methods 0.000 claims 5
- 238000007599 discharging Methods 0.000 claims 1
- 239000007921 spray Substances 0.000 claims 1
- 230000035699 permeability Effects 0.000 abstract description 4
- 238000005516 engineering process Methods 0.000 abstract description 2
- 238000013467 fragmentation Methods 0.000 abstract 1
- 238000006062 fragmentation reaction Methods 0.000 abstract 1
- 230000000737 periodic effect Effects 0.000 description 6
- 239000002360 explosive Substances 0.000 description 5
- 239000000843 powder Substances 0.000 description 5
- 230000008569 process Effects 0.000 description 5
- 238000000605 extraction Methods 0.000 description 4
- 238000002347 injection Methods 0.000 description 4
- 239000007924 injection Substances 0.000 description 4
- 239000003034 coal gas Substances 0.000 description 2
- 238000004519 manufacturing process Methods 0.000 description 2
- 238000009991 scouring Methods 0.000 description 2
- 238000005507 spraying Methods 0.000 description 2
- 230000009471 action Effects 0.000 description 1
- 230000002457 bidirectional effect Effects 0.000 description 1
- 230000015572 biosynthetic process Effects 0.000 description 1
- 238000004364 calculation method Methods 0.000 description 1
- 230000007547 defect Effects 0.000 description 1
- 238000010586 diagram Methods 0.000 description 1
- 238000009533 lab test Methods 0.000 description 1
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- E—FIXED CONSTRUCTIONS
- E21—EARTH OR ROCK DRILLING; MINING
- E21C—MINING OR QUARRYING
- E21C41/00—Methods of underground or surface mining; Layouts therefor
- E21C41/16—Methods of underground mining; Layouts therefor
- E21C41/18—Methods of underground mining; Layouts therefor for brown or hard coal
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- E—FIXED CONSTRUCTIONS
- E21—EARTH OR ROCK DRILLING; MINING
- E21B—EARTH OR ROCK DRILLING; OBTAINING OIL, GAS, WATER, SOLUBLE OR MELTABLE MATERIALS OR A SLURRY OF MINERALS FROM WELLS
- E21B33/00—Sealing or packing boreholes or wells
- E21B33/10—Sealing or packing boreholes or wells in the borehole
- E21B33/13—Methods or devices for cementing, for plugging holes, crevices or the like
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- E—FIXED CONSTRUCTIONS
- E21—EARTH OR ROCK DRILLING; MINING
- E21B—EARTH OR ROCK DRILLING; OBTAINING OIL, GAS, WATER, SOLUBLE OR MELTABLE MATERIALS OR A SLURRY OF MINERALS FROM WELLS
- E21B43/00—Methods or apparatus for obtaining oil, gas, water, soluble or meltable materials or a slurry of minerals from wells
- E21B43/006—Production of coal-bed methane
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- E—FIXED CONSTRUCTIONS
- E21—EARTH OR ROCK DRILLING; MINING
- E21B—EARTH OR ROCK DRILLING; OBTAINING OIL, GAS, WATER, SOLUBLE OR MELTABLE MATERIALS OR A SLURRY OF MINERALS FROM WELLS
- E21B43/00—Methods or apparatus for obtaining oil, gas, water, soluble or meltable materials or a slurry of minerals from wells
- E21B43/25—Methods for stimulating production
- E21B43/26—Methods for stimulating production by forming crevices or fractures
- E21B43/263—Methods for stimulating production by forming crevices or fractures using explosives
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- E—FIXED CONSTRUCTIONS
- E21—EARTH OR ROCK DRILLING; MINING
- E21B—EARTH OR ROCK DRILLING; OBTAINING OIL, GAS, WATER, SOLUBLE OR MELTABLE MATERIALS OR A SLURRY OF MINERALS FROM WELLS
- E21B7/00—Special methods or apparatus for drilling
- E21B7/04—Directional drilling
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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
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- F—MECHANICAL ENGINEERING; LIGHTING; HEATING; WEAPONS; BLASTING
- F42—AMMUNITION; BLASTING
- F42D—BLASTING
- F42D3/00—Particular applications of blasting techniques
- F42D3/04—Particular applications of blasting techniques for rock blasting
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- Environmental & Geological Engineering (AREA)
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Abstract
一种用于保护层开采坚硬顶板切顶方法，属于采矿工程矿压控制与切顶卸压技术领域，本发明采用理论计算爆破切顶位置，确定钻孔参数，开始钻孔施工、装药、封孔；按照理论结果进行阶段性爆破切顶；最后检验切顶效果。本发明以经典板结构理论、弹性基础梁理论为依据，结合实际顶板条件，采用爆破切顶的方法，缩短悬顶长度，有效控制了顶板垮落步距,操作性强、成本低、切顶速度快且效果好。尤其对于保护层开采，爆破切顶技术能有使煤岩体产生新的裂隙，扩大裂隙甚至贯通裂隙，增大煤岩体的破碎程度，极大地促进了瓦斯释放、流通，提高被保护层的透气性，进而提高抽采钻孔的预抽率，使瓦斯应力得以提前释放，有效地防治煤与瓦斯突出和冲击地压。The invention relates to a roof cutting method for hard roof mining for protective layer mining, belonging to the technical field of mine pressure control and roof cutting pressure relief in mining engineering. Seal the hole; carry out staged blasting and cutting according to the theoretical results; finally check the cutting effect. Based on the classical slab structure theory and the elastic foundation beam theory, the invention adopts the method of blasting and cutting the roof in combination with the actual roof conditions, shortens the length of the suspended roof, effectively controls the caving step of the roof, has strong operability, low cost, and cuts the roof. Fast and effective. Especially for protective layer mining, blasting and roof cutting technology can create new fissures in the coal and rock mass, expand the fissures or even penetrate the fissures, increase the degree of fragmentation of the coal and rock mass, greatly promote the release and circulation of gas, and improve the protection of the protected layer. The gas permeability can be improved, and the pre-drainage rate of the drainage hole can be improved, so that the gas stress can be released in advance, and the coal and gas outburst and rock burst can be effectively prevented.
Description
Technical Field
The invention belongs to the technical field of mining engineering mine pressure control and roof cutting pressure relief, and particularly relates to a forced roof cutting method for blasting a large-area exposed hard roof in a goaf by arranging bidirectional drill holes in roadways on two sides of a working face into the hard roof and placing special blasting cartridges in the drill holes.
Background
After stoping of the working face, the top plate is periodically collapsed under the action of mine pressure to form a working face goaf. However, a hard top plate with high strength, large thickness, strong integrity, no joint crack development and difficult natural collapse in a short period exists on the coal seam, so that the top plate of the goaf cannot fall along with the top plate of the working face to form a long-distance suspended roof state.
The long-distance suspended ceiling is easy to cause large steps of initial pressure and periodic pressure, dynamic pressure impact is generated during pressure, the support is damaged, and the typical impact of the top plate can cause the cylinder body of the upright post of the high-tonnage hydraulic support to explode; if the suspended roof falls, the generated strong storm can destroy and blow down the working face equipment, and casualties are caused. Particularly for the mining of a protective layer, the permeability coefficient of a coal seam is easily reduced because a large-area hard top plate is suspended without collapsing, so that coal gas is accumulated, the gas pressure is increased and exceeds the standard, and great difficulty is brought to the mining of the protective layer.
In order to solve the above problems, technical measures such as dismantling anchor cable nets and softening top plates by high-pressure water injection are generally adopted. But has the following disadvantages: firstly, an anchor cable net removing and adding process increases the cost; secondly, the roof plate still has the problem of long-distance suspended ceiling after the anchor is withdrawn; thirdly, high-pressure water injection has requirements on lithology, and the effect is not obvious on sandstone, conglomerate and other roof rocks which are not easily softened by water; fourthly, the collapse size of the top plate is not easy to control by high-pressure water injection; high pressure water is easy to mud cracks and joints to block the release of gas.
Disclosure of Invention
Aiming at the defects of the prior art, the invention provides a simple and feasible top cutting method which is easy to control and has good top cutting effect and can blast a hard top plate by reasonably arranging drill holes, so as to solve the problems of complicated working procedures, high production cost, unobvious effect, difficulty in control, long-distance suspended top formation and the like in the process of implementing technical measures such as anchor cable net disassembly, high-pressure water injection and the like.
The invention provides a hard roof cutting method for mining a protective layer, which comprises the following operation steps:
L1-first pressure step, m;
h is the thickness of the old roof rock stratum, m;
Rt-old top ultimate tensile strength, MPa;
q-load applied to the old roof, KN/m2；
L2＝(0.25～0.5)L1
L2-periodic incoming pressure step, m;
in order to ensure that the blasting and roof cutting effects are obvious, respectively arranging a drill site for roof cutting operation in two crossroads on a stope face at a distance 2 times of the initial pressure step distance from a cutting hole, setting the arranged drill sites as a first group, and setting the distance between each group of drill sites and the previous group of drill sites to be 2-3 times of the periodic pressure step distance;
3 drill holes are arranged in each drill site, namely an end cutting hole, an old top cutting hole and a block size control hole, wherein the hole pitch of the 3 drill holes is 1m, the 3 drill holes are arranged in an equilateral triangle, and the hole diameters are 60-75 mm;
wherein, the horizontal corner of the end cutting hole is 75 degrees, the elevation angle is 55 degrees, and the hole depth is 20 m; the horizontal corner of the old top cutting hole is 45 degrees, the elevation angle is 20 degrees to 27 degrees, and the hole depth is 45m to 60 m; the horizontal rotation angle of the block degree control hole is 0 degree, the elevation angle is 10 degrees to 15 degrees, and the hole depth is 60m to 85 m;
step 4, drilling construction and charging;
constructing in a drilling field according to the drilling parameters in the step 3, removing broken slag in the drilling hole by adopting a compressed air slag removal mode, immediately probing the hole by using a hole probing pipe after the blast hole construction is finished, verifying the hole depth, and immediately charging after the hole inspection is finished and is qualified in order to prevent hole collapse; wherein, the charge mode is: the blasting powder tube special for coal mine gas drainage is adopted for charging, the diameter of a charged powder bag is 32-45 mm, the length of the charged powder bag in an end cutting hole is 15m, the length of the charged powder bag in an old top cutting hole is 30-45 m, and the length of the charged powder bag in a block degree control hole is 42-60 m;
after drilling construction and charging, hole sealing operation is carried out; the hole sealing material is quick-setting expansion and non-shrinkage cement slurry, the hole sealing length of the end cutting hole is 5m, the hole sealing length of the old top cutting hole is 15m, and the hole sealing length of the block control hole is 18 m-25 m;
after hole sealing operation is finished, detonating a group of drill holes every 10 days, before detonating, sequentially checking a detonating network, checking line connection, detonating after detecting that the safety detonating standard is met, and spraying with wind and water during blasting, scouring rock slopes after blasting and opening a water curtain;
step 7, checking the effect of the top cutting operation;
monitoring and observing the stress and shrinkage of the hydraulic support on the top-cutting working surface to obtain the mine pressure display characteristic and the roof pressure coming law; and the influence of the top cutting operation on the mining of the protective layer is obtained by comparing and analyzing the gas content, the gas pressure of the coal body, the extraction concentration change and the change of the gas exhaust amount before and after the top cutting of blasting.
The hard roof topping method for overburden mining comprises the following steps:
in the step 6, the group of drill holes comprises 6 drill holes; from step 2, dividing the drill sites into groups, wherein each group comprises 2 drill sites, and then arranging 3 drill holes for each drill site in step 3, so that one group of drill holes comprises 6 drill holes; in addition, because the explosive connecting line is in a series connection mode between holes, the sequential inspection detonating circuit in the step is to inspect the detonating circuit along the series lead.
The invention has the following advantages:
according to the geological conditions of the working surface of the hard top plate, the blasting and roof cutting positions are calculated theoretically, drilling parameters are determined, and drilling construction, charging and hole sealing are started; performing staged blasting topping according to a theoretical result; and finally, checking the top cutting effect. The invention takes the classical plate structure theory and the elastic foundation beam theory as the basis, combines the actual roof condition, adopts the blasting roof cutting method, shortens the length of the suspended roof, effectively controls the roof caving step, and has strong operability, low cost, high roof cutting speed and good effect. Especially for the mining of a protective layer, the blasting roof cutting technology can enable a coal rock body to generate new cracks, expand the cracks and even penetrate the cracks, increase the crushing degree of the coal rock body, greatly promote the release and the circulation of gas, improve the air permeability of the protected layer, further improve the pre-pumping rate of an extraction drill hole, release the gas stress in advance, and effectively prevent and treat coal and gas outburst and rock burst.
Drawings
FIG. 1 is a front truncated borehole plan view of a working surface.
FIG. 2 is a front truncated borehole profile view of the working face.
FIG. 3 is a schematic view of a borehole placement within a drill site; in the figure 123, 3 different drilling sites are shown.
FIG. 4 is a schematic diagram of hole filling and sealing.
Detailed Description
The process of the present invention will be further illustrated with reference to examples.
The mining depth of a working face of a certain ore protective layer is 1100m, the trend is as long as 572m, the trend is as long as 150m, and a trend long-wall total caving coal mining method and a retreat type stoping method are selected. The mining face support adopts a hydraulic support to control a top plate, the maximum top control distance is 5.14m, and the minimum top control distance is 4.54 m. According to geological drilling data, the thickness of the coal seam is 1.0m, the inclination angle of the coal seam is 6 degrees, and the coal seam belongs to a thin coal seam. The direct roof is 0-6 m of sandy mudstone, the old roof is 9m of medium-grain sandstone, and the roof belongs to a hard roof. There are problems in that: when mining by a caving coal mining method, the hard top plate can be suspended in a large area in a goaf without caving, dynamic pressure impact is generated during pressure application, and the support is easy to damage; because the large-area roof is hung without caving, a large amount of gas is accumulated in a goaf, the gas concentration is seriously overrun, and the mine production is seriously damaged.
A hard roof cutting method for mining a protective layer comprises the following specific operation steps:
the old top is medium sandstone with a thickness of 9.0m, and the ultimate tensile strength R of the medium sandstone is obtained through laboratory testst6.5MPa, and the load q of the overlying coal seam on the old top is 3.6 multiplied by 10 according to a coal seam histogram5From this, the step L of the first coming pressure of the old crown is obtained1Is 54m, the periodic pressure step L2The maximum is 27 m.
two crossroads on a stope face are respectively provided with a drill site for top cutting operation at a position which is 2 times of the initial pressure step distance from a cutting hole, namely 108m, the arranged drill sites are set as a first group, the distance between each group of drill sites and the previous group of drill sites is 70m, namely about 2.6 times of the period pressure step distance, and the specific arrangement is shown in figure 1.
3 drill holes are distributed in each drill site, namely an end cutting hole, an old top cutting hole and a block size control hole, as shown in figure 3, the hole spacing of the 3 drill holes is 1m and are distributed in an equilateral triangle, the diameter of the upper air inlet lane drill hole is 60mm, and the diameter of the lower air inlet lane drill hole is 75 mm;
as shown in fig. 1 and 2, the horizontal corner of the end cutting hole is set to be 75 degrees, the elevation angle of the upper air inlet lane is set to be 55 degrees, the hole depth is set to be 20m, the elevation angle of the lower air inlet lane is set to be 55 degrees, and the hole depth is set to be 20 m; the horizontal corner of the old top cutting hole is 45 degrees, the elevation angle of the upper air inlet lane is 20 degrees, the hole depth is 45m, the elevation angle of the lower air inlet lane is 27 degrees, and the hole depth is 60 m; the horizontal rotation angle of the block degree control hole is 0 degree, the elevation angle of the upper air inlet lane is 10 degrees, the hole depth is 60m, the elevation angle of the lower air inlet lane is 15 degrees, and the hole depth is 85 m.
Step 4, drilling construction and charging;
and (4) constructing in the drilling field according to the drilling parameters in the step (3), and removing slag in the drilled hole by adopting a compressed air slag removal mode so as to ensure that the subsequent hole exploring tube and the medicine tube can be smoothly transferred. After the blast hole construction is finished, immediately probing the hole by using a hole probing pipe, and verifying the depth of the hole, wherein the hole is verified to be finished and qualified;
wherein, the charge mode is: the explosive charging of the blasting explosive tube special for coal mine gas drainage is adopted, the diameter of the explosive charge loaded in the upper air inlet lane is 32mm, and the diameter of the explosive charge loaded in the lower air inlet lane is 45 mm. The length of the medicine bag filled in the end cutting hole is 15m, the length of the medicine bag filled in the upper air inlet lane old top cutting hole is 30m, the length of the medicine bag filled in the lower air inlet lane old top cutting hole is 45m, the length of the medicine bag filled in the upper air inlet lane block size control hole is 42m, and the length of the medicine bag filled in the lower air inlet lane block size control hole is 60 m.
after drilling construction and charging, hole sealing operation is carried out as shown in fig. 4; the hole sealing material is quick-setting expansion and non-shrinkage cement paste, and the length of the end cutting hole is 5 m; the sealing length of the old top cut-off hole is 15 m; the hole sealing length of the upper air inlet roadway block-size control hole is 18m, and the hole sealing length of the lower air inlet roadway block-size control hole is 25 m.
and after hole sealing operation is finished, detonating a group of drill holes every 10 days, before detonating, sequentially checking a detonating network, checking line connection, detonating after detecting that the safety detonating standard is met, and spraying by using wind and water, scouring rock walls after blasting and opening a water curtain during blasting.
Step 7, monitoring and observing stress and shrinkage of the hydraulic support on the opposite-roof-cutting working surface; and comparing and analyzing parameters such as coal gas pressure, extraction concentration change and the like before and after blasting and roof cutting.
The results show that roof collapse is characterized by: when the working face is mined to the position of 52m, the middle part of the bracket begins to collapse sporadically. And (4) the mining face is pushed to 58m, the first group of top cutting holes are blasted, and most of hard top plates collapse after blasting. The maximum pressure of the bracket is 26MPa, and the step distance of the pressure of the old top in the first time can be inferred to be 55m, which is basically consistent with the theoretical calculation. The working face is pushed to 83m, and the comprehensive collapse occurs: most of the stent collapsed to the forced caving area. From the above roof collapse process, it can be known that: the initial pressure step pitch should be 55m, and the periodic pressure step pitch should be 20 m-30 m. In the process of collapse, the pressure of the support is not obvious, generally between 26 and 27MPa, and is within a normal controllable range.
Gas ginsengThe number variation is characterized by: before and after the blasting of the return airway, the average pumping concentration in the range of 30m before and after the blasting position is increased from 6.5 percent to 57 percent, the average pumping negative pressure is reduced from 236mmHg to 230mmHg, and the average differential pressure is increased by 0.2mmH2O, average daily allowance from 641m3Is raised to 742m3(ii) a The average pumping concentration in the range of 30m before and after the blasting position of the air inlet roadway is increased from 18 percent to 38.8 percent, the average pumping negative pressure is reduced from 235mmHg to 210mmHg, and the average pressure difference is reduced by 0.4mmH2O, average daily drainage is 984m3Rise to 1251m3。
In conclusion, the roof caving step is effectively controlled by the protective layer hard roof mining roof cutting method, and the method is strong in operability, high in roof cutting speed and remarkable in effect. Meanwhile, the coal body cracks are enlarged and even run through the cracks, the breaking degree of the coal rock body is increased, the gas release and circulation are greatly promoted, the air permeability of a protected layer is improved, the pre-pumping rate of the extraction drill hole is further improved, and a good effect is achieved.
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Cited By (10)
| Publication number | Priority date | Publication date | Assignee | Title |
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| CN113218261A (en) * | 2021-04-15 | 2021-08-06 | 华能煤炭技术研究有限公司 | Top plate blasting and top breaking method |
| CN113216977A (en) * | 2021-04-15 | 2021-08-06 | 华能煤炭技术研究有限公司 | Blasting method for rear broken top of working face frame |
| CN114060088A (en) * | 2021-11-19 | 2022-02-18 | 安徽理工大学 | A kind of prevention method of layer separation water damage of coal seam roof in mine |
| CN114278293A (en) * | 2021-12-03 | 2022-04-05 | 天地科技股份有限公司 | Method for cutting top of coal seam roof |
| CN115046443A (en) * | 2022-06-14 | 2022-09-13 | 安徽理工大学 | Construction method for breaking blasting in top-cut and bottom-cut of thick coal seam roof-cutting roadway retaining |
| CN115822600A (en) * | 2022-09-09 | 2023-03-21 | 贵州大学 | Methods of controlling non-uniform roof caving in hard rock formations |
| CN116104474A (en) * | 2023-02-17 | 2023-05-12 | 安徽纬培矿山科技有限责任公司 | Roof cutting pressure relief method based on roof cutting drilling hole edge concentrated stress fracturing principle |
| CN117948143A (en) * | 2024-03-26 | 2024-04-30 | 煤炭科学技术研究院有限公司 | Roof fracturing method for coal mine working face in initial mining stage |
| CN118390996A (en) * | 2024-03-18 | 2024-07-26 | 中国矿业大学 | A method for efficiently extracting gas from coal seams with increased permeability in coal tunnel top cut |
| CN119825363A (en) * | 2025-03-20 | 2025-04-15 | 中煤科工开采研究院有限公司 | Methods for improving top coal caving performance |
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Cited By (13)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN113218261A (en) * | 2021-04-15 | 2021-08-06 | 华能煤炭技术研究有限公司 | Top plate blasting and top breaking method |
| CN113216977A (en) * | 2021-04-15 | 2021-08-06 | 华能煤炭技术研究有限公司 | Blasting method for rear broken top of working face frame |
| CN114060088A (en) * | 2021-11-19 | 2022-02-18 | 安徽理工大学 | A kind of prevention method of layer separation water damage of coal seam roof in mine |
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