Decompression and Anti-Impact method by 3D Stratified Buffer Energy-Absorbing Belt for Thick coal seam

The present invention belongs to the technical field of preventing and ground pressure impact-controlling in mining industry, specifically applied in the bottom road of a thick coal seam in a coal mine, it relates to a method of unloading pressure and preventing impact in a thick coal seam by an energy-absorbing belt with three-dimensional layered buffer.

As the scale and depth of coal mining increase year by year, the intensity and harm of ground pressure shocking impact become more and more prominent. Ground pressure impact disaster occurs in the process of mining, destroying the mining face and roadway and other mining systems, resulting in a large number of equipment damage, casualties and economic losses. After a lot of theoretical research and engineering practice on the mechanism and technical of preventing and controlling ground pressure impact, the industry have successively formed the technical solutions of preventing and controlling ground pressure impact, such as water injection in the coal seam, unloading pressure by large-diameter drilling, unloading by blasting in the coal seam, deep-hole blasting in the top plate, and fracturing in the top plate of the coal seam etc. However, large-diameter drill holes usually have a small range of pressure relief, result in a heavy work load, and there are also other issues, such as the fracturing of the top plate of the coal seam is disorderly and the range of fracturing is not controllable. Further more the blasting of the coal seam and the top and bottom plate usually being affected by the control of mining pyrotechnic materials, which makes it difficult to source the materials and result in high cost. Thus, there are limitations of the above technological measures in the field of the prevention and control of ground pressure impact applications.

According to the previous research and substantial engineering practices, taking effective pressure relief measures in the coal seam to reduce the stress concentration in the coal seam is the main prevention and control principle for ground pressure impact prevention and control. However, due to the complex geological conditions of thick coal seams, and the affection of the top and bottom plate stresses of coal seams, lateral stresses of roof plate in the mining area, supporting pressure of the mining face, and layout of the working face, the prevention and control measures against ground pressure impact are insufficient and uneven, which cannot effectively reduce the risk of impact, which significantly threat the safety of production in the mines. Thick coal seams with gas hazards are usually arranged with bottom plate roadway, and the construction of drilling holes through the seams is carried out to unload the pressure. With the development of water jet technology and equipment, cutting slots in the coal seam to form a pressure relief buffer energy absorption zone can provide pressure relief deformation space to reduce the elastic energy accumulation, blocking stress transfer, reducing stress concentration, to effectively reducing the risk of impact pressure in the coal seam. However, there is no mature theoretical system as the basis for the implementation of water jet pressure unloading and impact prevention in thick coal seams, and the parameters of on-site water jet cutting coal body rely on on-site engineering experience to formulate the implementation plan, which is uncontrollable, and may easily result in insufficient unloading of the coal body in the mined area. Therefore, the prevention and control of pressure impact in coal seams is not ideal under existing technologies.

In view of this, an object of the present application is to provide a method of unloading pressure and preventing impact in a thick coal seam by a three-dimensional layered buffer and energy-absorbing belt, to solve the problem of insufficient and uneven unloading of pressure in thick coal seams of impact ground pressure mines.

In order to achieve the above objectives, the present invention provides the following technical solutions:

A method for unloading pressure and preventing shocking impact in a three-dimensional stratified buffer energy-absorbing belt of a thick coal seam, characterized in that it comprises the following steps:

Optionally, in step S1, when 0≤σi≤1.5 σc and there is an incremental increase in stress within the range from the front of the working face, it is determined that this interval is the length of the stress increase zone range L∈[Lσc, L1.5σc]; when σi≥1.5 σc within the range from the front of the working face, this interval is determined that this interval is the length of the stress critical load zone range L=L≥1.5σc; when the front face of the work and the stress is decreasing, the interval is determined as the length of the stress drop zone L∈[L1.5σc, Lσc].

Optionally, in step S1, the range of stress static loading area Lis derived according to the evolution law of support stress distribution in the working face, L=K1 (L+L+L), K1 is the work imbalance coefficient, and K1 is taken as 1.2˜1.5.

Optionally, in step S2, a row of stress prediction boreholes is constructed at intervals of 30 m in the direction of advancement of back-mining face, and at least 3 boreholes are constructed in each row, controlling the two sides of the trench and the central position respectively.

Optionally, in step S2, along the advancing direction of the working face, the spacing between the rows of cut-and-sew drilling holes within the range of stress-rising area and stress-decreasing area is 3 m; the spacing between the rows of cut-and-sew drilling holes within the range of stress-critical load area is 2 m; and the spacing between the rows of cut-and-sew drilling holes within the range of stress-quiet load area is 4 m.

Optionally, in step S3, the statistical stress prediction borehole takes an average value of the debris output per unit depth of coal penetration, Wi, and a value of the change in the debris output, ΔW.

Optionally, in step S3, when the value of the change in the amount of chips out ΔW∈[−0.2, 0.2] Kg/m3, it will be the height of the stress stabilisation zone H; when the value of the change in the amount of chips out ΔW∈[0.2, 0.5] Kg/m3, it will be the height of the stress stabilisation zone H; and when ΔW≥0.5 Kg/m3, it will be the height of the stress stacking zone H.

Optionally, in step S4, along the vertical direction of the working face, the spacing of the borehole cuttings is 3 m within the height of the coal body stress stabilisation zone H, the spacing of the borehole cuttings is 2 m within the height of the coal body stress increase zone H, and the spacing of the borehole cuttings is 1 m within the height of the coal body stress superposition zone H.

Optionally, in step S4, supplementary slit drilling is constructed to cut coherent barrier slots at the height of the stress stabilisation zone of the coal body H, the height of the stress augmentation zone of the coal body H, and the height of the stress superposition zone of the coal body H, respectively.

Optionally, water jetting anti-hedging and pressure relief construction is carried out in the coal seam bottom lane for the anti-hedging unit.

The beneficial effect of the present invention is that: based on the stress distribution law in the advancing direction and vertical direction of the back-mining face of the thick coal seam, the present invention implements water jet pressure relief drilling in the stress concentration area of the back-mining area in the coal seam bottom lane, and artificially creates a buffer-absorbing and pressure-relieving anti-punching belt inside the coal seam. Different jet cutting parameters are adopted in different stress areas to cut the coal body, so as to effectively reduce the stress concentration inside the coal body. At the same time, according to the basis of stress division of the thick coal seam, cut continuously at the junction of different divisions within the coal body to form a blocking groove, which not only plays the role of pressure relief of the coal body, but also can block the transmission of stress. By dividing the stress distribution of the thick coal seam, adopting reasonable jet parameters in different stress areas, and adopting the method of unloading pressure in layers inside the coal seam to solve the problems of insufficient and uneven unloading of the coal seam, the three-dimensional unloading of the thick coal seam has been formed, which effectively reduces the danger of pressure impact in the coal seam.

Other advantages, objects and features of the present invention will to some extent be set forth in the subsequent specification and to some extent will be apparent to those skilled in the art based on an examination of the following or can be taught from the practice of the invention. The objects and other advantages of the present invention may be achieved and obtained by the following specification.

In order to make the objects, technical solutions and advantages of the present invention clearer, the invention will be described in preferred detail below in conjunction with the accompanying drawings, in which:

shows the schematic diagram of the distribution of the support stress curve along the advancing direction of the working face in the thick coal seam;

shows the schematic diagram of drilling construction arrangement along the working face advance direction for thick coal seam;

shows the construction diagram of vertical section of thick coal seam three- dimensional unloading and anti-punching

Attachment labels: coal seam, backing face-, coal seam unloading area-, coal seam floor, bottom drawdown lane, working face support stress curve, cuttings borehole, stress prediction borehole, length of stress elevation zone L, length of stress critical loading zone L, length of stress drop zone L, length of stress static loading zone L, roadway stress curve, superimposed stress curve of the working face, cuttings slot, blocking slots, Supplementary slit drilling, Stress stabilisation zone height H, Stress increase zone height H, Stress superposition zone height H.

The following illustrates the embodiments of the present invention by means of particular specific examples, and other advantages and efficacies of the present invention can be readily understood by those skilled in the art from the contents disclosed in this specification. The present invention may also be implemented or applied in different specific embodiments, and the details in this specification may be modified or changed in various ways based on different views and applications without departing from the spirit of the present invention. It is to be noted that the illustrations provided in the following embodiments illustrate the basic concept of the present invention in a schematic manner only, and the following embodiments and the features in the embodiments may be combined with each other without conflict.

Among them, the accompanying drawings are for exemplary illustration only, and represent only schematic drawings, not physical drawings, and cannot be understood as a limitation of the present invention; in order to better illustrate the embodiments of the present invention, certain parts of the accompanying drawings will be omitted, enlarged, or reduced, and do not represent the dimensions of the actual product; it is understandable to the person skilled in the art that certain well-known structures and their descriptions in the accompanying drawings may be omitted.

In the accompanying drawings of the embodiments of the present invention, the same or similar symbols correspond to the same or similar parts; in the description of the present invention, it is to be understood that the terms “up”, “down”, “left”, “right”, “front”, “back” and the like indicate orientation or positional relationships based on the orientation or positional relationships shown in the accompanying drawings only for the purpose of illustration. “up”, “down”, “left”, “right”, “front”, “back” and the like indicate orientation or positional relationships based on those shown in the accompanying drawings, and are used only for the purpose of facilitating the description of the present invention and simplifying the description. The description of the present invention and the simplification of the description, rather than indicating or implying that the device or element referred to must have a specific orientation, be constructed and operated in a specific orientation, therefore the terms describing the positional relationships in the accompanying drawings are only used for exemplary illustration and are not to be construed as a limitation of the present invention, and the specific meanings of the above terms may be understood by a person of ordinary skill in the field in accordance with the specific circumstances.

Referring to, a three-dimensional layered buffer energy-absorbing belt pressure-unloading and anti-impact method for thick coal seams is based on the stress distribution law in the advancing direction and the vertical direction of the back-mining face of thick coal seams, and water jet pressure-unloading drilling is carried out in the bottom lane of the coal seams in the stress-concentrated area of the back-mining area, and the buffer energy-absorbing belt is formed by cutting the barrier grooves inside the coal seams. Firstly, the distribution length L of the supporting stress curve is estimated along the advancing direction of the working face, and an independent anti-shock unit is divided. Determine the length of the stress increase zone L, the length of the stress critical load zone L, the length of the stress decrease zone L, and the length of the stress static load zone L, and determine different spacing parameters of the cutter drill holes in different stress zones. In the working vertical direction, determine the height of coal body stress stabilisation zone H, the height of stress increase zone H, the height of stress superposition zone H, and determine the parameters of the slit and the location of the blocking groove in different stress zones. Reasonable jet parameters are adopted in different stress areas, and barrier slots are formed by continuous cutting at the junction of different partitions, which not only play the role of coal unloading, but also block the stress transfer, solve the problems of insufficient and uneven unloading of coal seams, and form the three-dimensional unloading of thick seams, which effectively reduces the risk of impact pressure in coal seams.

This includes the following steps:

S2 According to the supporting stress distribution law in the advancing direction of the back-mining face, determine the spacing of the arrangement of slit drill holesand stress prediction drill holesin different subdivisions of a three-dimensional independent unit for the prevention and control of ground pressure impact.

S3 Stress prediction drilling adopts the statistical method of drilling chip volume to deduce the stress distribution pattern in the vertical direction of the working face, and reasonably classify the height of the stress increase zone Hand the height of the stress superposition zone H.

S4 According to the law of stress distribution in vertical direction of thick coal seam mining back in the face, determine the parameters of the cut seam and the location of the blocking groove in different stress zones.

S5 With reference to the construction parameters determined in steps S1˜S4, carry out the construction of the anti-impact pressure unloading works on the remaining anti-impact units to form a three-dimensional pressure unloading space in the entire coal seam.

Preferably, the length L of the distribution field of the support stress curve of the working face is determined by the law of distribution of the support stress in front of the 1-1 front of the return mining face, i.e. L=L+L+L L+. The coal body within the length L of the distribution of the support stress curve of the working face in front of the working face of the thick seam is treated as a three-dimensionally independent unit of the prevention and control of the impact ground pressure.

Preferably, based on the theory of Rock Control, the strength of the coal body σc is calculated based on the burial depth of the coal seam, which can be measured and calculated by means of CT stress scanning, stress gauges, numerical simulation analyses and other means of calculating the working face support stress σi at Li in front of the 1-1 front of the return mining face.

Preferably, when 0≤σi≤1.5 σc and there is an increasing stress in the range from the front of the workpiece, the interval is determined to be a stress elevation zone length range L∈[Lσc, L1.5σc].

Preferably, the interval is determined to be the stress critical load zone length range L=L≥1.5σc when σi≥1.5σc from the front of the working face range.

Preferably, the interval is determined to be a stress drop zone length range L∈[L1.5σc, Lσc] when 1.5σc≥σi≥σc and there is a decreasing stress condition from the front side of the work.

Preferably, when σi=σc in the range from the front of the working face, it can be determined that this interval is the range of the stress static load zone L. Based on the evolution law of the work face support stress distribution, L=K1 (L+L+L), K1 is the work imbalance coefficient, and K1 is generally taken to be 1.2˜1.5.

Preferably, the purpose of constructing the stress prediction drill holes is to obtain the stress distribution law in the vertical direction of the working face, and to provide a basis for determining the parameters of the spacing of the cuttings for the cuttings drill holes. A row of stress prediction drill holes is constructed at intervals of 30 m from the advancing direction of the return mining face, and each row of drill holes is at least 3, respectively controlling the working two sides of the trough and the central position.

Preferably, the slit boreholesare spaced at different parameters within different stress zones. The spacing of the slit drilling holesis 3 m within the stress rise and stress fall zones, 2 m within the stress critical load zone, and 4 m within the stress static load zone.

Preferably, the chip-out per unit depth of coal penetration from the coal sighting point to the final hole point is counted for each of the stress prediction drill holesW1i, W2i, W3i, i being the unit depth of coal penetration. The chip output per unit depth of coal penetration for the three stress prediction drill holesis averaged Wi=(W1i+W2i+W3i)/3.

Preferably, the value of the change in the amount of chips per unit depth of the coal seam ΔW=Wi+1−Wi. When ΔW∈[−.,.] Kg/m3, reflecting a more uniform distribution of stresses within the scope of the region of the depth of penetration of the coal can be determined as the height of the stress stabilisation zone H.

Preferably, the value of the change in chip output per unit depth of the coal seam ΔW=Wi+1−Wi. When ΔW∈[0.2, 0.5] Kg/m3, the chip output within the range of this penetration depth region gradually increases, reflecting the gradual increase in the stress of the coal body, and the height of the stress stabilisation zone Hcan be determined.

Preferably, the change value of chip-out per unit depth of the coal seam is calculated as ΔW=Wi+1−Wi. When ΔW≥0.5 Kg/m3, the chip-out within the range of this penetration depth region is obviously increased, reflecting that the interior of the coal body has been extruded due to the high stress, and the height of the stress superimposed zone can be determined H

Preferably, the drilling cuttings are spaced vertically 3 m apart within the height Hof the coal body stress stabilisation zone, 2 m apart vertically within the height Hof the coal body stress augmentation zone, and 1 m apart vertically within the height Hof the coal body stress superposition zone.

When the barrier slotscannot be made to pass through by the slit drilling holes, coherent barrier slotscan be cut by constructing supplementary slit drilling holesat the heights of the stress stabilisation zone of the coal body H, the height of the stress stabilisation zone of the coal body H, and the height of the stress superposition zone of the coal body H, respectively, so as to form a cushioning and absorbing zone between the barrier slots.

This embodiment takes the coal second layer of an impact ground pressure mine as a geological background, which has the danger of protruding and impacting compound power disaster. The back-mining face is located in the southern part of the sixth mining area of the second seam of coal, with the ground elevation between 1995 and 2320 m, and the burial depth between 594 and 777 m, and the mined coal seam is the second seam of coal, with the thickness of 25.04-45.0 m and the average of 34.5 m, and the mining height of the working face is 5 m, and the inclination angle of the coal seam is 5°-15°, and the working face is designed to have the strike length of 584 m, the inclination width of 120 m, and the strength of the coal body is σc=18 MPa.

A method of unloading pressure and preventing impact in a thick coal seam with three-dimensional stratified buffer and energy absorbing belt, referring to, which show a coal seam, a backing face-, a coal seam unloading area-, a coal seam footing, a bottom drawdown lane, a working face support stress curve, a cuttings drilling hole, a stress prediction drilling hole, a length of a stress elevation zone L, a length of a stress critical loading zone L, a length of a stress decrease zone L, a length of a stress static loading zone L, Tunnel stress curve, working face superimposed stress curve, slit groove, barrier groove, supplementary slit borehole, stress stabilisation zone height H, stress increase zone height H, stress superimposed zone height H, comprising the steps of:

S1 According to the basic geological data, such as the conditions of the coal seam, the characteristic parameters of coal body mechanics, and the layout of the working face, the CT stress scanning method was used to study the stress distribution law along the advancing direction of the working face. In the range of 0˜25 m in front of the working face, the coal body stress σ is 0˜27 MPa with an increasing trend, in the range of 25˜40 m, the stress σis more than 27 MPa, and the maximum value is 40 MPa, which shows the trend of increasing and then decreasing, in the range of 40˜70 m, the stress σ is 18˜27 MPa, and the stress σ is about 18 MPa after the range of 70 m.

According to the above stress distribution range, it can be seen that the stress in front of the working face shows the state of increasing and then decreasing and tends to moderate, and the composite working face support stress distribution law. According to the division of stress rise area (0≤σi≤1.50σc), stress critical load area (σi≥1.50σc), stress drop area (1.5σc≥σi≥σc), it can be determined that the length of the stress rise area is 35 m, the length of the stress critical load area is 40 m, and the length of the stress drop area is 70 m in turn.

According to the working face supporting stress distribution evolution law, L=K1 (L+L+L23), K1 is the work imbalance coefficient, K1 take 1.5, can calculate the length of the stress static load zone L=1.5×(25+15+30)=105 m; Determine the length of the distribution field of the supporting stress curve of the working face, i.e., L=L+L+L+L=25+15+30+105=175 m; Therefore, the coal seam within the range of 175 m in front of the working face can be treated as a three-dimensional independent unit; the three-dimensional independent unit to prevent impact ground pressure is divided by the length of working face according to the management of a separate unit. Therefore, the coal seam within the range of 175 m in front of the working face can be regarded as a three-dimensional independent unit for preventing and controlling impact ground pressure; the three-dimensional independent unit for preventing and controlling impact ground pressure is divided by the length of the working face, and those that don't satisfy the length of an independent unit are managed according to a separate unit. Therefore, the length of working faceis 584 m, n=[1/L]+1=[584÷175]+1=4, so the working face can be divided into 4 units.