Prevention and control method

Abstract

The invention provides a prevention and control method which comprises the following steps: performing roof presplitting blasting on a thick and hard rock stratum at a higher position of an overlying rock stratum, so that a roof is regularly broken; wherein the single-layer thickness of the thick and hard rock stratum is larger than or equal to a preset thickness value, and the uniaxial compressive strength of the thick and hard rock stratum is larger than or equal to a preset strength value; and the top plate of the lower layer of the overlying rock layer is subjected to blasting breakage increasing, so that the goaf space is filled with caved rocks.

Description

Technical Field

[0001] The present invention relates to the technical field of coal mine safety, and in particular to a prevention and control method. Background Art

[0002] With the continuous increase in the depth of coal resource mining, rock burst has gradually become one of the major disasters threatening coal mine safety production. In addition, the mining conditions of rock burst mines are complex, and most accidents occur in the form of multiple disasters coupled.

[0003] The geological conditions of rock burst mines are complex, and the existing complex disaster prevention and control technologies cannot effectively solve the problems encountered in mine production.

[0004] For example, a roof weakening method pre-treats the hard roof through hydraulic fracturing measures, modifies and weakens the hard roof, promotes the development of internal cracks in the strip weakening, reduces its overall strength, and thus maintains the safety of the tunnels below the hard roof.

[0005] For another example, a method of retaining lanes along the goaf in deep wells with small coal pillars based on unloading the roof of the key layer below the roof, in which blasting is carried out on the side of the coal seam roof that is biased towards the solid coal. In the key layer below the pre-cracked coal seam roof, the stress transfer of the key layer is cut off, and the high stress acting on the small coal pillars and the lanes retained along the goaf is transferred, thereby improving the lane environment and facilitating the maintenance of the lane.

[0006] However, the existing technical solutions only crush and weaken the high-level hard roof, so as to achieve the transfer of high stress and improve the stress environment of the tunnel surrounding rock; such technical solutions are only applicable to situations where the geological conditions are simple, the roof is weak in nature, and there is no aquifer in the overlying rock strata. If the geological conditions are more complicated, such as the roof is relatively hard, only crushing and weakening the high-level hard roof can still easily lead to a large area of ​​hanging roof; and if the low-level rock strata are relatively hard, it is difficult to fully fill the space of the goaf after collapse, resulting in a higher height of the collapsed rock strata required to fill the goaf, causing the water-conducting fracture zone to develop into the aquifer and increase the risk of mine water inrush.

[0007] Existing technical solutions usually adopt the roof pre-splitting blasting method, which is a common measure for preventing and controlling rock burst in coal mines. However, after blasting the hard roof, the impermeable layer above the coal seam is destroyed, which will increase the risk of mine water inrush. Summary of the invention

[0008] The main purpose of the present invention is to provide a prevention and control method to simultaneously solve the problems of rock burst and water inrush.

[0009] In order to achieve the above-mentioned purpose, the present invention provides a prevention and control method, which includes: carrying out roof pre-splitting blasting on the thick and hard rock layer in the higher layer of the overlying rock layer to make the roof cross-broken regularly; wherein the thick and hard rock layer is a rock layer with a single layer thickness greater than or equal to a preset thickness value and a uniaxial compressive strength greater than or equal to a preset strength value; blasting and crushing the roof of the lower layer of the overlying rock layer to make the collapsed rock fill the space in the goaf.

[0010] Furthermore, the roof of the lower layer of the overlying rock layer is blasted and crushed to form a collapse zone; according to the formula Predict the development height of the collapse zone; where h 1 is the development height of the collapse zone, M is the mining thickness of the coal seam, k is the rock looseness coefficient, and a is the inclination of the coal seam.

[0011] Furthermore, the thick hard rock layer at a higher level of the overlying rock layer is subjected to roof pre-splitting blasting to form a fracture zone; according to the formula h 2 =C*h 1 Predict the development height of the fault zone; where h 2 is the development height of the fault zone, C is the coefficient, and the value range of C is 1 to 3.

[0012] Furthermore, the preset thickness value is 10m.

[0013] Furthermore, the preset strength value is 60Mpa.

[0014] Furthermore, before implementing roof pre-splitting blasting on the thick hard rock layers in the higher layers of the overlying strata and blasting and fragmenting the roof in the lower layers, the prevention and control methods also include: exploring the on-site geological conditions to determine the location of the aquifer in the overlying strata; and drilling and coring the roof above the tunnel to measure rock mechanical parameters to determine the location and thickness of the key layers in the overlying strata.

[0015] Furthermore, the blasting holes of the working face are arranged in a fan shape, with 3 or 4 blasting holes arranged in each fan-shaped section.

[0016] Furthermore, the diameter of the blasting hole is 75 mm.

[0017] Furthermore, the blasting holes are arranged near the shoulder of the tunnel.

[0018] Furthermore, the blast hole inclination angle is calculated according to the formula Calculate; where is the blast hole inclination angle, h is the height difference between the blast hole opening position and the final hole position, l s It is the horizontal distance between the opening position and the final position of the blasting hole.

[0019] Furthermore, the charge of a single blast hole is calculated according to the formula Q = q*l c Calculate; where Q is the charge, q is the charge density of the explosive line, l c is the charge length.

[0020] Furthermore, the charge amount of a single blasting hole is less than or equal to 100 kg.

[0021] Furthermore, the charge length of a single blasting hole is greater than or equal to 10m.

[0022] Furthermore, the sealing length of the blasting hole is greater than or equal to one third of the depth of the blasting hole.

[0023] Furthermore, the sealing length of the blasting hole is greater than or equal to 5m.

[0024] Applying the technical solution of the present invention, the prevention and control method includes: performing roof pre-splitting blasting on the thick hard rock layer at a higher level of the overlying rock layer to make the roof cross-broken regularly; wherein the thick hard rock layer is a rock layer with a single layer thickness greater than or equal to a preset thickness value and a uniaxial compressive strength greater than or equal to a preset strength value; blasting and fragmenting the roof of the lower level of the overlying rock layer to make the collapsed rock fill the space in the goaf.

[0025] During the specific implementation process, pre-splitting blasting is carried out on the thick and hard rock strata at higher levels of the overlying rock strata to promote regular roof fractures and roof cavitation as the mining progresses, avoiding large areas of roof hanging and reducing the risk of impact during the working face mining.

[0026] The blasting method is used to blast the roof of the lower layers of the overlying rock strata to increase its fragmentation and expandability. The collapsed rock can fill more space in the goaf, thereby reducing the height of the collapsed rock strata required to fill the unfilled goaf and reducing the height of the overlying rock damage. The height of the water-conducting fracture zone does not reach the aquifer, thereby achieving the purpose of preventing and controlling water inrush from the working face. BRIEF DESCRIPTION OF THE DRAWINGS

[0027] The drawings constituting a part of the present application are used to provide a further understanding of the present invention. The exemplary embodiments of the present invention and their descriptions are used to explain the present invention and do not constitute an improper limitation of the present invention. In the drawings:

[0028] Figure 1 A schematic diagram showing the arrangement of the high-position roof pre-splitting blasting holes according to the present invention is shown;

[0029] Figure 2 A schematic diagram of the arrangement of low-position roof fragmentation blasting holes according to the present invention is shown. DETAILED DESCRIPTION

[0030] It should be noted that, in the absence of conflict, the embodiments and features in the embodiments of the present application can be combined with each other. The present invention will be described in detail below with reference to the accompanying drawings and in combination with the embodiments.

[0031] It should be noted that the following detailed descriptions are illustrative and are intended to provide further explanation of the present application. Unless otherwise specified, all technical and scientific terms used herein have the same meanings as those commonly understood by those skilled in the art to which the present application belongs.

[0032] It should be noted that the terms used herein are only for describing specific embodiments and are not intended to limit the exemplary embodiments according to the present application. As used herein, unless the context clearly indicates otherwise, the singular form is also intended to include the plural form. In addition, it should be understood that when the terms "comprise" and / or "include" are used in this specification, it indicates the presence of features, steps, operations, devices, components and / or combinations thereof.

[0033] The present invention provides a prevention and control method, which is an integrated prevention and control method for composite geological disasters of rock burst and water inrush, including high-position roof breaking and low-position increased fragmentation. The method mainly includes two aspects, namely, high-position roof breaking and low-position roof increased fragmentation by blasting, so as to solve the problems of rock burst and water inrush at the same time.

[0034] Specifically, the prevention and control methods include: carrying out pre-splitting blasting on the roof of the thick and hard rock layers in the higher layers of the overlying rock strata to make the roof cross-broken regularly; wherein, the thick and hard rock strata are rock strata with a single layer thickness greater than or equal to a preset thickness value and a uniaxial compressive strength greater than or equal to a preset strength value; blasting and fragmenting the roof of the lower layers of the overlying rock strata to make the collapsed rock fill the space in the goaf.

[0035] During the specific implementation process, pre-splitting blasting is carried out on the thick and hard rock strata at higher levels of the overlying rock strata to promote regular roof fractures and roof cavitation as the mining progresses, avoiding large areas of roof hanging and reducing the risk of impact during the working face mining.

[0036] The blasting method is used to blast the roof of the lower layers of the overlying rock strata to increase its fragmentation and expandability. The collapsed rock can fill more space in the goaf, thereby reducing the height of the collapsed rock strata required to fill the unfilled goaf and reducing the height of the overlying rock damage. The height of the water-conducting fracture zone does not reach the aquifer, thereby achieving the purpose of preventing and controlling water inrush from the working face.

[0037] Optionally, the preset thickness value is 10m; the preset strength value is 60Mpa. That is, a rock layer with a single layer thickness exceeding 10m and a uniaxial compressive strength exceeding 60MPa is a thick hard rock layer, also called a hard rock layer.

[0038] Specifically, the overburden strata form collapse zones, fault zones, and bending zones; the fault zones are also called fissure zones. The roof of the lower layers of the overburden strata is blasted and crushed to form collapse zones; the roof of the thick and hard rock layers of the higher layers of the overburden strata is pre-splitting blasted to form fault zones.

[0039] Specifically, the development height of the collapse zone and the development height of the fault zone are both mainly related to factors such as the thickness of the mined coal seam, inclination, overburden lithology, coal mining methods, and roof management.

[0040] Specifically, according to the formula Predict the development height of the collapse zone; where h 1 is the development height of the collapse zone, in meters; M is the mining thickness of the coal seam, in meters; k is the rock looseness coefficient; a is the inclination of the coal seam, in degrees.

[0041] Specifically, according to the formula h 2 =C*h 1 Predict the development height of the fault zone; where h 2 is the development height of the fault zone, in meters; C is the coefficient, and the value range of C is 1 to 3.

[0042] In the present application, before the roof pre-splitting blasting is performed on the thick hard rock layer at the higher level of the overlying rock layer and the roof blasting and fragmentation is performed on the lower level, the prevention and control method also includes:

[0043] The geological conditions on site are explored and based on the geological exploration results, the location of aquifers in the overlying rock strata is determined.

[0044] Drill holes and take cores from the roof above the tunnel to measure the rock mechanics parameters; and based on the results of the rock mechanics parameter measurements, determine the position and thickness of the key layer in the overlying rock strata; among them, the rock strata with a single layer thickness greater than or equal to the preset thickness value and a uniaxial compressive strength greater than or equal to the preset strength value are thick hard rock strata.

[0045] Specifically, the geological exploration results include information such as rock strata lithology, layer thickness, cumulative burial depth, etc. That is, the rock strata lithology, layer thickness, cumulative burial depth, etc. are determined based on the borehole column chart.

[0046] Specifically, conventional hydrogeological drilling is used to lay out geological holes, mainly to understand the geological and hydrogeological conditions, such as stratigraphic sequence, lithology, thickness, fracture development, horizontal and vertical distribution of aquifers and impermeable layers.

[0047] Specifically, rock mechanics parameters include rock compressive strength, tensile strength, density, porosity, elastic modulus, Poisson's ratio, cohesion, internal friction angle and other information.

[0048] Specifically, the rock layer that controls the movement of the overlying rock layer in the mining area or all the rock layers on the surface is called the key layer, which has the following characteristics: 1. Thickness; 2. High strength; 3. When the key layer sinks and deforms, the sinking amount of the overlying rock layer is synchronized and coordinated; 4. The rupture of the key layer will trigger the synchronous rupture of the overlying rock layer; 5. The main body of the rock layer load.

[0049] In this application, the opening and final hole positions of the blasting holes shall be determined comprehensively based on the site conditions, the position of the key layers, the blasting rock formations, etc.

[0050] Optionally, the blasting holes are opened near the shoulder of the tunnel.

[0051] In the present application, the blasting holes of the working face are arranged in a fan shape, with 3 or 4 blasting holes arranged in each fan-shaped section. Figure 1 There are four blasting holes arranged in the left fan-shaped section, namely SK1, SK2, SK3, and SK4; Figure 1 There are three blasting holes arranged in the right fan-shaped section, namely SK5, SK6, and SK7; Figure 2 There are four blasting holes arranged in the left fan-shaped section, namely QK1, QK2, QK3, and QK4; Figure 2 There are four blasting holes arranged in the right fan-shaped section, namely QK5, QK6, QK7 and QK8.

[0052] In this application, the inclination angle of the blasting hole should be determined comprehensively based on the opening hole position, the final hole position, etc.

[0053] Specifically, the blast hole inclination angle is calculated according to the formula Calculate. Among them, is the blast hole inclination angle, in degrees; h is the height difference between the blast hole opening position and the final hole position, in meters; l s It is the horizontal distance between the blast hole opening position and the final hole position, in meters.

[0054] In the present application, the length of the blasting hole is determined comprehensively based on the hole opening position, the hole end position, etc.

[0055] Specifically, the blast hole penetrates the entire hard rock roof.

[0056] In the present application, optionally, the diameter of the blasting hole is 75 mm.

[0057] In this application, the charge amount for a single hole should be determined comprehensively based on the position, thickness, strength, etc. of the blasting rock layer.

[0058] Specifically, the single-hole charge is the charge of a single blasting hole; the single-hole charge is calculated according to the formula Q = q*l cCalculate. Where Q is the charge amount, in kilograms; q is the charge density of the explosive line, in kilograms / meter; l c is the charge length in meters.

[0059] Optionally, the single hole charge is less than or equal to 100 kg.

[0060] In the present application, the single-hole charging length refers to the length of a single blasting hole; the single-hole charging length is determined comprehensively based on the opening hole position and the final hole position.

[0061] Optionally, the single-hole charging length is greater than or equal to 10m.

[0062] For example, depending on the distance and inclination from the tunnel to the key layer, the single-hole charging length can be set to 10m, 12m, 15m, 17m, or 21m.

[0063] In the present application, the sealing length of the blasting hole is greater than or equal to one third of the depth of the blasting hole; and / or, the sealing length of the blasting hole is greater than or equal to 5 m.

[0064] like Figure 1 As shown, high-position roof pre-splitting blasting holes are constructed from 20m outside the cutting eyes of the two drifts of the working face (respectively the transport drift and the return air drift) to the stop-mining line, with a hole spacing of 10m; 4 drill holes (i.e. blasting holes) are arranged in each group of transport drifts, namely SK1, SK2, SK3, and SK4; 3 drill holes (i.e. blasting holes) are arranged in each group of return air drifts, namely SK5, SK6, and SK7.

[0065] The elevation angle of SK1 is 90 degrees, the azimuth angle is 0 degrees, the total length of the hole is 40m, the hole diameter is 75mm, the charging length is 10m, the charging amount is 30Kg, and the sealing length is 30m.

[0066] The elevation angle of SK2 is 75 degrees, the azimuth angle is 30 degrees, the total length of the hole is 41m, the hole diameter is 75mm, the charging length is 10m, the charge is 31Kg, and the sealing length is 31m.

[0067] The elevation angle of SK3 is 55 degrees, the azimuth angle is 330 degrees, the total length of the hole is 49m, the hole diameter is 75mm, the charging length is 12m, the charge is 36Kg, and the sealing length is 37m.

[0068] The elevation angle of SK4 is 35 degrees, the azimuth angle is 0 degrees, the total length of the hole is 70m, the hole diameter is 75mm, the charging length is 17m, the charge is 52Kg, and the sealing length is 52m.

[0069] The elevation angle of SK5 is 90 degrees, the azimuth angle is 0 degrees, the total hole length is 40m, the hole diameter is 75mm, the charging length is 10m, the charge is 30Kg, and the sealing length is 30m.

[0070] The elevation angle of SK6 is 75 degrees, the azimuth angle is 30 degrees, the total hole length is 41m, the hole diameter is 75mm, the charging length is 10m, the charge is 31Kg, and the sealing length is 31m.

[0071] The elevation angle of SK7 is 60 degrees, the azimuth angle is 330 degrees, the total hole length is 46m, the hole diameter is 75mm, the charging length is 12m, the charge is 34Kg, and the sealing length is 35m.

[0072] It should be noted that the total hole length is also the hole depth.

[0073] like Figure 2 As shown, low-position roof crushing blasting holes are constructed from 20m outside the cutting eyes of the two drifts of the working face (respectively the transport drift and the return air drift) to the stop-mining line; in order to increase the crushing and expansion of the low-position roof rock strata, the hole spacing of the low-position crushing blasting holes is 5m; 4 low-position crushing holes are arranged in each group of transport drifts, namely QK1, QK2, QK3, and QK4; 4 low-position crushing holes are arranged in each group of return air drifts, namely QK5, QK6, QK7, and QK8.

[0074] The elevation angle of QK1 is 90 degrees, the azimuth angle is 0 degrees, the total length of the hole is 30m, the hole diameter is 75mm, the charging length is 12m, the charging amount is 35Kg, and the sealing length is 18m.

[0075] The elevation angle of QK2 is 75 degrees, the azimuth angle is 30 degrees, the total length of the hole is 31m, the hole diameter is 75mm, the charging length is 12m, the charge is 37Kg, and the sealing length is 19m.

[0076] The elevation angle of QK3 is 55 degrees, the azimuth angle is 330 degrees, the total hole length is 37m, the hole diameter is 75mm, the charging length is 15m, the charge is 43Kg, and the sealing length is 22m.

[0077] The elevation angle of QK4 is 35 degrees, the azimuth angle is 0 degrees, the total length of the hole is 52m, the hole diameter is 75mm, the charging length is 21m, the charge is 62Kg, and the sealing length is 31m.

[0078] The elevation angle of QK5 is 90 degrees, the azimuth angle is 0 degrees, the total hole length is 30m, the hole diameter is 75mm, the charging length is 12m, the charge is 35Kg, and the sealing length is 18m.

[0079] The elevation angle of QK6 is 75 degrees, the azimuth angle is 30 degrees, the total length of the hole is 31m, the hole diameter is 75mm, the charging length is 12m, the charge is 37Kg, and the sealing length is 19m.

[0080] The elevation angle of QK7 is 60 degrees, the azimuth angle is 330 degrees, the total hole length is 35m, the hole diameter is 75mm, the charging length is 14m, the charge is 41Kg, and the sealing length is 21m.

[0081] The elevation angle of QK8 is 35 degrees, the azimuth angle is 0 degrees, the total length of the hole is 52m, the hole diameter is 75mm, the charging length is 21m, the charge is 62Kg, and the sealing length is 31m.

[0082] From the above description, it can be seen that the above embodiments of the present invention achieve the following technical effects:

[0083] The purpose of this application is to prevent and control the combined disasters of rock burst and water inrush in rock burst mines rich in water layers.

[0084] The prevention and control methods include: carrying out pre-splitting blasting on the roof of the thick and hard rock layers in the higher layers of the overlying rock strata so that the roof can be regularly broken; wherein, the thick and hard rock strata are rock strata with a single layer thickness greater than or equal to a preset thickness value and a uniaxial compressive strength greater than or equal to a preset strength value; blasting and fragmenting the roof of the lower layers of the overlying rock strata so that the collapsed rock can fill the space in the goaf.

[0085] The composite geological disaster prevention and control method of high-position roof breaking and low-position fragmentation provided in the present application can realize the roof pre-splitting blasting of high-position thick and hard rock layers, so as to promote the roof to burst as the mining progresses, avoid large-area roof hanging, and reduce the impact risk during the mining process of the working face; it can also realize the blasting and fragmentation of low-position rock layers, improve their fragmentation and expansion properties, so that the collapsed rock layers can fill more goaf space, so as to reduce the height of the collapsed rock layers required for filling the unfilled goaf, and avoid the development of water-conducting fracture zones into aquifers and increase the risk of water inrush in mines; while adopting roof pre-splitting blasting to prevent and control impact ground pressure, it can also reduce the height of overburden damage and prevent water inrush in mines.

[0086] It should be noted that the terms "first", "second", etc. in the specification and claims of the present application and the above-mentioned drawings are used to distinguish similar objects, and are not necessarily used to describe a specific order or sequence. It should be understood that the data used in this way can be interchangeable where appropriate, so that the embodiments of the present application described herein can be implemented in an order other than those illustrated or described herein, for example. In addition, the terms "including" and "having" and any of their variations are intended to cover non-exclusive inclusions, for example, a process, method, system, product or device that includes a series of steps or units is not necessarily limited to those steps or units that are clearly listed, but may include other steps or units that are not clearly listed or inherent to these processes, methods, products or devices.

[0087] For ease of description, spatially relative terms such as "above", "above", "on the upper surface of", "above", etc. may be used here to describe the spatial positional relationship between a device or feature and other devices or features as shown in the figure. It should be understood that spatially relative terms are intended to include different orientations of the device in use or operation in addition to the orientation described in the figure. For example, if the device in the accompanying drawings is inverted, the device described as "above other devices or structures" or "above other devices or structures" will be positioned as "below other devices or structures" or "below other devices or structures". Thus, the exemplary term "above" can include both "above" and "below". The device can also be positioned in other different ways (rotated 90 degrees or in other orientations), and the spatially relative descriptions used here are interpreted accordingly.

[0088] The above description is only a preferred embodiment of the present invention and is not intended to limit the present invention. For those skilled in the art, the present invention may have various modifications and variations. Any modification, equivalent replacement, improvement, etc. made within the spirit and principle of the present invention shall be included in the protection scope of the present invention.

Claims

1. A prevention and control method, characterized in that: include: Carry out roof pre-splitting blasting on thick hard rock layers at higher levels of the overlying rock layers so as to cause the roof to be regularly broken; wherein the thick hard rock layers are rock layers with a single layer thickness greater than or equal to a preset thickness value and a uniaxial compressive strength greater than or equal to a preset strength value; The roof of the lower layers of the overlying rock strata is blasted and broken up so that the collapsed rock can fill the space in the goaf.

2. The control method according to claim 1, characterized in that: The roof of the lower layer of the overlying rock layer is blasted and broken to form a collapse zone; According to the formula The development height of the collapse zone is predicted; wherein h1 is the development height of the collapse zone, M is the mining thickness of the coal seam, k is the rock looseness coefficient, and a is the inclination of the coal seam.

3. The prevention and control method according to claim 2, characterized in that: Pre-splitting blasting of the roof of thick and hard rock layers at higher levels of the overburden is carried out to form a fracture zone; The development height of the fault zone is predicted according to the formula h2=C*h1; wherein h2 is the development height of the fault zone, C is a coefficient, and the value range of C is 1 to 3.

4. The control method according to claim 1, characterized in that: The preset thickness value is 10m; and / or The preset strength value is 60Mpa.

5. The prevention and control method according to claim 1, characterized in that: Before carrying out roof pre-splitting blasting on thick hard rock layers at higher levels of the overlying rock formations and blasting and fragmenting the roof layers at lower levels, the control method further comprises: The on-site geological conditions are explored to determine the location of the aquifer in the overburden; and cores are drilled from the roof above the tunnel to measure rock mechanical parameters to determine the location and thickness of the key layers in the overburden.

6. The control method according to claim 1, characterized in that: The blasting holes of the working face are arranged in a fan-shaped pattern, with 3 or 4 blasting holes arranged in each fan-shaped section; and / or The blast hole diameter is 75mm; and / or The blasting holes are located near the shoulder of the tunnel.

7. The control method according to claim 1, characterized in that: The blast hole inclination angle is calculated according to the formula Calculate; where is the blast hole inclination angle, h is the height difference between the blast hole opening position and the final hole position, l s It is the horizontal distance between the opening position and the final position of the blasting hole.

8. The control method according to claim 1, characterized in that: The charge of a single blast hole is calculated according to the formula Q = q*l c Calculate; where Q is the charge, q is the charge density of the explosive line, l c is the charge length.

9. The control method according to claim 1 or 8, characterized in that: The charge of a single blast hole is less than or equal to 100 kg; and / or The charging length of a single blasting hole is greater than or equal to 10m.

10. The control method according to claim 1, characterized in that: The length of the blast hole sealed is greater than or equal to one third of the blast hole depth; and / or The sealing length of the blasting hole is greater than or equal to 5m.

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