CN112610251B - Control method of coal mining roadway top plate - Google Patents
Control method of coal mining roadway top plate Download PDFInfo
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- CN112610251B CN112610251B CN202011527124.7A CN202011527124A CN112610251B CN 112610251 B CN112610251 B CN 112610251B CN 202011527124 A CN202011527124 A CN 202011527124A CN 112610251 B CN112610251 B CN 112610251B
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- E—FIXED CONSTRUCTIONS
- E21—EARTH DRILLING; MINING
- E21D—SHAFTS; TUNNELS; GALLERIES; LARGE UNDERGROUND CHAMBERS
- E21D17/00—Caps for supporting mine roofs
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- E—FIXED CONSTRUCTIONS
- E21—EARTH DRILLING; MINING
- E21D—SHAFTS; TUNNELS; GALLERIES; LARGE UNDERGROUND CHAMBERS
- E21D20/00—Setting anchoring-bolts
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- E—FIXED CONSTRUCTIONS
- E21—EARTH DRILLING; MINING
- E21F—SAFETY DEVICES, TRANSPORT, FILLING-UP, RESCUE, VENTILATION, OR DRAINING IN OR OF MINES OR TUNNELS
- E21F15/00—Methods or devices for placing filling-up materials in underground workings
Abstract
The disclosure relates to a control method of a coal mining roadway roof, which comprises the following steps: determining the height of a joint according to the coal mining height, the roadway height and the initial crushing expansion coefficient of a roof rock mass; and forming a top plate kerf on the roadway top plate according to the determined kerf height. According to the control method of the coal mining roadway roof, the initial crushing expansion coefficient of the roof rock mass is considered in the design parameters of the actual upper seam height, the seam cutting height is determined according to the coal mining height, the roadway height and the initial crushing expansion coefficient of the roof rock mass, the caving gangue can be filled in a goaf completely after crushing expansion, the roof of the high level can be contacted with the gangue once sinking, and then the gangue is compacted continuously until the gangue is stable.
Description
Technical Field
The disclosure relates to the technical field of mining industry, in particular to a control method of a coal mining roadway roof.
Background
The longwall mining method was first introduced in the united kingdom, schlopshire, early in the eighteenth century, and was subsequently used in coal mines in other counties, staford, debarkshire, etc. In the 30 th century in China, experiments are firstly carried out on a certain coal mine in Shandong province, and through development of nearly one hundred years, a mature longwall mining theory, technology and equipment system is formed at present and is widely applied to mines in all parts of the country. When the method is adopted for mining, 2 stoping roadways need to be tunneled in advance when 1 working face is mined, and 1 roadway protection coal pillar is reserved between adjacent working faces. This mining method results in the failure to recover a large amount of coal resources and high tunneling costs.
The gob-side entry retaining technology is always a hot topic of domestic and foreign research, and the technology can solve the problems of coal resource waste and large tunnel excavation investment to a certain extent. At present, the most common gob-side entry retaining mode is roadside support by adopting an artificial filling body. However, the problems of stress concentration of the filling body, mismatching of filling and mining efficiency and the like are often difficult to solve by adopting the mode.
In order to solve the problems of stress concentration of a filling body, mismatching of filling and mining efficiency and the like, an N00 construction method is provided, namely N working faces in a mining area are mined without tunneling a roadway in advance and reserving coal pillars. The N00 method cannot be applied directly because the conventional roof control method does not take into account the characteristic structural features of the roof.
It is to be noted that the information disclosed in the above background section is only for enhancement of understanding of the background of the present disclosure, and thus may include information that does not constitute prior art known to those of ordinary skill in the art.
Disclosure of Invention
The purpose of the present disclosure is to overcome the deficiencies of the prior art, and provide a control method for a coal mining roadway roof, which can fill a goaf completely just after the crushed and expanded caving rock, and once the high-level roof sinks, the high-level roof contacts with the caving rock, and then continuously compacts the caving rock until the situation is stable.
According to an aspect of the present disclosure, there is provided a control method of a coal mining roadway roof, the control method including:
determining the height of a joint according to the coal mining height, the roadway height and the initial crushing expansion coefficient of a roof rock mass;
and forming a top plate kerf on the roadway top plate according to the determined kerf height.
In an exemplary embodiment of the present disclosure, the kerf height is:
wherein HfIs the height of the cutting seam HmFor coal mining height, HrIs the height of the roadway, K0Is the initial crushing and swelling coefficient of the roof rock mass.
In an exemplary embodiment of the present disclosure, the control method further includes:
determining constant-resistance anchor cable supporting force of the top plate before the short-arm beam is broken according to anchor cable supporting force required before the arm beam is broken, roadway width, joint cutting height, joint cutting angle, distance between acting point of anchor cable supporting resultant force and roadway entity coal side, dead weight of the short-arm beam structure, and supporting force of collapsed waste rock on the short-arm beam structure in the vertical direction;
determining constant-resistance anchor cable supporting force of the broken short arm beam of the top plate according to anchor cable supporting force required after the arm beam is broken, the self weight of the short arm beam structure, the distance between the position of a broken point of the short arm beam structure and an entity coal side, the position of the entity coal side, supporting force of collapsed waste rocks on the short arm beam structure in the vertical direction, and the distance between the action point of the combined anchor cable supporting force and the entity coal side of the roadway;
and according to the determined constant-resistance anchor cable supporting force required before and after the short-arm beam is broken, supporting the top plate within the range of the seam by using the constant-resistance anchor cable.
In an exemplary embodiment of the present disclosure, the required constant-resistance anchor cable supporting force before the short-arm beam breaks is:
wherein, Fc1The constant-resistance anchor cable supporting force required before the arm beam is broken is a roadway widthDegree HfIs the height of the cutting seam, beta is the angle of the cutting seam, LcThe distance between the acting point of the resultant force of the anchor cable support and the solid coal wall of the roadway, G is the self weight of the short-arm beam structure, and F is the distance between the acting point of the resultant force of the anchor cable support and the solid coal wall of the roadwaygThe support force of the short-arm beam structure in the vertical direction is realized for the collapse waste rock.
In an exemplary embodiment of the present disclosure, the required constant-resistance anchor cable supporting force after the short-arm beam is broken is:
wherein, Fc2The constant-resistance anchor cable supporting force required after the arm beam is broken is G, the dead weight of the short arm beam structure is x0The distance between the position of the short-arm beam structure fracture point and the solid coal side, qcFor the position of the solid coal side, FgSupporting force of the short-armed beam structure in the vertical direction for caving waste rock, LcThe distance between the acting point of the combined force of the anchor cable support and the entity coal side of the roadway is obtained.
In an exemplary embodiment of the present disclosure, the control method further includes:
determining dynamic pressure bearing temporary supporting force according to dynamic pressure bearing temporary supporting force, roadway width, average volume weight of overlying strata, volume weight of a short-arm beam structure, joint cutting height, joint cutting angle, support coefficient of caving gangue to a short-arm beam, stress concentration coefficient caused by mining, coal seam burial depth, number of each row of constant-resistance anchor cables, internal friction angle of the short-arm beam structure, cohesion of the short-arm beam structure, pretightening force of a single constant-resistance anchor cable and supporting force of the caving gangue to the short-arm beam structure in the vertical direction;
and carrying out dynamic pressure bearing support on the top plate according to the determined dynamic pressure bearing temporary support force.
In one exemplary embodiment of the present disclosure, characterized in that,
wherein, FtTo moveBearing temporary support force, a is the width of the roadway, gamma1Is the average volume weight of overburden, gamma2Is the volume weight of the short-armed beam structure, HfIs the height of the cutting seam, beta is the angle of the cutting seam, K1Is the support coefficient of the collapsed waste rock to the short arm beam, K2H is the coal seam burial depth, n is the number of each row of constant-resistance anchor cables,
is the internal friction angle of the short-armed beam structure, c is the cohesion of the short-armed beam structure, fpPre-tightening force for single constant-resistance anchor cable, FgThe support force of the short-arm beam structure in the vertical direction is realized for the collapse waste rock.
In an exemplary embodiment of the present disclosure, the determined constant-resistance anchor rope supporting force required before and after the short-arm beam is broken is multiplied by a safety factor to form the required constant-resistance anchor rope supporting force.
In an exemplary embodiment of the present disclosure, the determined dynamic pressure bearing temporary support force is multiplied by a safety factor to form a required dynamic pressure bearing temporary support force.
In an exemplary embodiment of the present disclosure, the safety factor is 1.5-2.
According to the control method of the coal mining roadway roof, the initial crushing expansion coefficient of the roof rock mass is considered in the design parameters of the actual upper seam height, and the changed crushing expansion coefficient of the gangue in the goaf gangue compaction process is not taken as the design basis. The change process of the residual crushing expansion coefficient and the crushing expansion coefficient influences the final deformation amount and the deformation process of the pressure-bearing structure, and the change process has no direct relation with the height of the cutting seam. Therefore, the height of the joint cutting is determined through the coal mining height, the roadway height and the initial crushing expansion coefficient of the top plate rock mass, the goaf can be filled completely just after the crushed expansion of the caving gangue, the top plate of the high-rise position is contacted with the gangue once sinking, and then the gangue is compacted continuously until the gangue is stable.
It is to be understood that both the foregoing general description and the following detailed description are exemplary and explanatory only and are not restrictive of the disclosure.
Drawings
The accompanying drawings, which are incorporated in and constitute a part of this specification, illustrate embodiments consistent with the present disclosure and together with the description, serve to explain the principles of the disclosure. It is to be understood that the drawings in the following description are merely exemplary of the disclosure, and that other drawings may be derived from those drawings by one of ordinary skill in the art without the exercise of inventive faculty.
Fig. 1 is a flowchart of a control method for a coal mining roadway roof according to an embodiment of the present disclosure.
Fig. 2 is a flowchart of a control method for a coal mining roadway roof according to another embodiment of the present disclosure.
Fig. 3 is a flowchart of a control method for a coal mining roadway roof according to still another embodiment of the present disclosure.
Fig. 4 is an N00 construction method mining face and roadway layout provided for one embodiment of the present disclosure.
Fig. 5A is a schematic view of a shearer head roadway provided in an embodiment of the disclosure.
Fig. 5B is a schematic view of a constant-resistance anchor cable reinforcement according to an embodiment of the present disclosure.
Fig. 5C is a schematic view of a top plate orientation slit provided by an embodiment of the present disclosure.
Fig. 5D is a schematic view of a dynamic pressure bearing support provided by an embodiment of the present disclosure.
Fig. 5E is a schematic illustration of goaf roof caving provided in an embodiment of the present disclosure.
Fig. 5F is a schematic illustration of a stabilization zone reinforcement support provided by one embodiment of the present disclosure.
Fig. 6 is a schematic structural view of a top plate of a roadway according to an embodiment of the present disclosure.
Fig. 7A is a schematic design diagram of a roof kerf height when a coal mining height is smaller than a roadway height according to an embodiment of the present disclosure.
Fig. 7B is a schematic design diagram of a roof kerf height with a coal mining height greater than a roadway height according to an embodiment of the disclosure.
Fig. 8A is a schematic diagram illustrating the design strength of the constant-resistance anchor cable support before the short-armed beam structure breaks according to an embodiment of the present disclosure.
Fig. 8B is a schematic diagram illustrating the design strength of the constant-resistance anchor cable support after the arm-beam structure is broken according to an embodiment of the disclosure.
Fig. 9 is a schematic view of a dynamic pressure bearing support design according to an embodiment of the present disclosure.
Fig. 10 is a schematic diagram of a test work surface provided by an embodiment of the present disclosure.
Fig. 11 is a graph illustrating the resultant force of anchor cable support and the relationship between the resultant force and the distance from the solid coal roadway side according to an embodiment of the present disclosure.
Fig. 12 is a schematic diagram of a constant resistance anchor cable and roof lancing parameter design provided in an embodiment of the present disclosure.
Fig. 13 is a schematic view of dynamic pressure bearing support parameters according to an embodiment of the present disclosure.
Fig. 14 is a schematic application effect diagram provided by an embodiment of the disclosure.
Detailed Description
Example embodiments will now be described more fully with reference to the accompanying drawings. Example embodiments may, however, be embodied in many different forms and should not be construed as limited to the embodiments set forth herein; rather, these embodiments are provided so that this disclosure will be thorough and complete, and will fully convey the concept of example embodiments to those skilled in the art. The same reference numerals in the drawings denote the same or similar structures, and thus their detailed description will be omitted. Furthermore, the drawings are merely schematic illustrations of the present disclosure and are not necessarily drawn to scale.
The terms "a," "an," "the," "said," and "at least one" are used to indicate the presence of one or more elements/components/parts/etc.; the terms "comprising" and "having" are intended to be inclusive and mean that there may be additional elements other than the listed elements/acts.
In order to solve the problems of stress concentration of a filling body, mismatching of filling and mining efficiency and the like, a 110 construction method is provided, namely, only 1 roadway needs to be tunneled in advance and 0 coal pillar is reserved for each mining 1 working face. In the latter, an N00 construction method is further provided, namely N working faces in one mining area are mined without the need of tunneling a roadway in advance and reserving coal pillars. The two methods reduce the amount of roadway tunneling work and the amount of coal pillar reservation, are beneficial to weakening stress concentration caused by traditional mining, and can bring remarkable economic benefit and safety benefit for coal mines. The two mining methods are similar in the control principle of the roof structure, namely the caving height of the roof of the goaf is enlarged by using the roof cutting, and the subsidence of the overlying strata is controlled by using the collapsed waste rock after crushing and expansion, so that the stability of the roadway is ensured. Therefore, the roadway formed by the two methods is also called a cut-top gob-side entry retaining method.
Compared with the traditional longwall mining method, the roof structure under the mining condition of the N00 construction method has the following unique characteristics: 1) a special short-arm beam structure is formed on a top plate without a coal pillar self-lane under the effect of a joint cutting; 2) one side of the roadway is of a scattered body structure consisting of caving waste rocks, and the caving waste rocks are one of main bearing bodies of the top plate of the high-rise position; 3) the roadway is formed behind the working face, and is not influenced by the dynamic pressure of the driving roadway and the advance stress of the working face, so that the stability of the roadway is better. The N00 construction method does not need to consider roadway control ahead of the working face. All roadway range top plates needing to be controlled are in short-arm beam structures instead of combined beam structures with two fixed ends. This is the main difference between the N00 method and the 110 method. The conventional support theory and design method do not consider the characteristics, so that the method cannot be directly applied to the roadway support, and a new support design method needs to be searched.
When the N00 method is adopted for coal mining, a roadway (except a first mining face) which is tunneled in advance is not needed in front of a working face, and the roadway is gradually formed behind the working face in the coal mining process. The N00 method has the unique advantages that more than 80% of tunneling engineering amount in a mining area can be eliminated, and no coal pillar needs to be reserved. Therefore, the extraction rate of the coal can be greatly improved, the production cost is obviously reduced, and the problem of continuous tension of excavation can be thoroughly solved. Note that the N00 method referred to herein is the first generation technology. With this method, a circle of boundary roadways around the panel are to be tunneled. The second generation technology can realize boundary laneways without tunneling, and is currently carrying out engineering tests on the Xintai coal mine in Yanan City of China. The arrangement of the working face and the roadway of the N00 engineering method is shown in figure 4.
In order to achieve the above object, firstly, the problem of cutting coal by a coal mining machine under the condition that no roadway is arranged in front of a working face is solved, the space of the roadway is cut by the coal mining machine, and then the space is reserved as the roadway by adopting a series of supporting technologies, and the specific process is as follows:
1) by modifying the matching mode of the coal mining machine, the scraper conveyor and the working face support, the coal mining machine can surpass the scraper conveyor to cut coal, and then a tunnel space and an arc-shaped tunnel side wall are cut at the end of the working face by using the roller, as shown in fig. 5A.
2) After the coal mining machine cuts out the roadway space, the top plate of the space is reinforced by using the constant-resistance large-deformation anchor cable, as shown in fig. 5B. The constant-resistance large-deformation anchor cable has excellent mechanical characteristics such as high prestress and high constant resistance, and can suspend the top plate in a certain range on a stable rock stratum of a high level, so that the deformation of the top plate of a low level is effectively controlled by means of the stability of the top plate of the high level.
3) The physical connection between the goaf roof and the roadway roof is cut off by utilizing the roof directional joint cutting technology, so that the roof in a certain range of the roadway forms a short-arm beam structure, the roofs on the two sides of the joint cutting have mutually independent motion characteristics, the influence of goaf gangue collapse on the roadway roof is eliminated, and the stress state of the roadway roof is improved, as shown in fig. 5C.
4) The dynamic pressure bearing support device closely follows the tail of the working face hydraulic support, utilizes a dynamic pressure bearing support technology to temporarily support a top plate, and resists the dynamic pressure influence on a roadway in the movement process of the top plate at a high position. The dynamic pressure bearing and supporting structure must have the characteristics of high initial supporting force, rapid resistance increase, high working resistance and the like, and also have certain retractility, as shown in fig. 5D.
5) After the roof is subjected to joint cutting and supporting, the working face support moves forwards, and the goaf roof near the roadway quickly collapses under the action of the joint cutting to form a bulk roadway side. Meanwhile, the collapsed rock mass fills the mining space under the crushing and expanding effect, and forms a certain degree of support for the top plate at the high position, so that the movement space after the fracture is reduced, and the new balance structure is formed quickly, as shown in fig. 5E.
6) After the low-level roof of the goaf is collapsed, the high-level roof is also fractured and generates rotary sinking. When the collapsed waste rock is compacted, the sinking of the top plate of the high-rise position gradually stops. But here the sink stop is not absolute. Over time, the caving gangue and roof also undergo slow creep deformation. Therefore, for the roadway with longer service life, the reinforcing support of the stable area should be carried out, as shown in fig. 5F.
The mechanism of structural stability of the top plate: in a certain range behind the working face, although the waste rock in the goaf collapses, the initial accumulation state of the waste rock is usually looser, the loose waste rock is further compressed under the action of the pressure of the overlying strata, and a high-level hard top plate above the joint seam is still fractured and rotationally sunk, so that a strong dynamic pressure influence is generated on the roadway. When the collapsed gangue is compacted, the gangue supporting force and the coal body supporting force can enable the fractured roof to form new balance, the entry is kept in a stable state, and the stable roof structure is shown in fig. 6.
As can be seen from fig. 6, the roof structure affecting the stability of the roadway in the course of the pillar-free self-entry mining mainly has two parts, the first part is a high-rise hard roof above the cutting seam and is called as a pressure-bearing structure; the second part is a top plate in the kerf range and is called a short-arm beam structure. The two top plate structures respectively have different deformation characteristics, so different control ideas are adopted.
For the pressure bearing structure, a rock A above the coal body, a rock B above the roadway and a rock C above the goaf form a 'masonry beam' or 'transmission rock beam' bearing structure together. One end of the rock block B is fractured in the coal body, and generates rotary sinking by taking the fracture position as the center, and the other end of the rock block B is contacted with the collapsed waste rock and gradually compacted after sinking, so that a balance structure with one end supported by the solid coal body and the other end supported by the waste rock is finally formed. The structure is a main bearing body for overburden pressure, and the motion state of the structure plays a decisive role in the stability of a lower roadway. According to prior studies, the rotational deformation generated by the structure is generally regarded as "given deformation", which is difficult to prevent by in-lane bracing. To achieve control of the bearing structure, the bearing capacity of the rock itself must also be used. Therefore, the roof cutting mode is proposed to control roof collapse within a certain range, and then the broken expanded gangue after collapse is fully utilized to support the roof at the high level, so that the roof is enabled to form a self-stabilizing structure.
For a short-arm beam structure, one end of the short-arm beam structure is connected with the internal stable rock stratum, and the other end of the short-arm beam structure is separated from the goaf top plate along the seam cutting surface. Under the effect of roof joint-cutting, the influence of goaf waste rock caving on the short-arm beam structure is obviously weakened. After the top plate above the cutting seam is stable, the structure can keep balance under the combined action of coal body supporting force, roadway supporting resistance and the like under the shield of broken rock blocks of the high-level hard top plate. Therefore, aiming at the top plate in the range, a high-strength and high-prestress supporting mode is adopted to anchor the short-arm beam structure to the hard top plate at the high-rise position, so that the bearing capacity of the hard top plate at the high-rise position is fully utilized to realize the stability of the roadway.
The most problems existing at present are that the research on the design criteria and the quantitative design method of the technologies is not deep enough, and how to take values of a plurality of key design parameters is still unclear. Therefore, the comprehensive design method and the calculation basis for the roof control are provided by combining the stability characteristics of the roof of the gob-side entry retaining of the N00 construction method.
1) The bearing structure is a main bearing body for overburden pressure, and the bearing capacity and the motion state of the bearing structure directly influence the stability of a roadway below the bearing structure. Whether accurate roof cutting can be carried out becomes the most critical link for controlling the stability of the surrounding rock. On one hand, the accurate roof cutting aims to cut the roof within a certain range by utilizing the roof cutting seam, so that the roadway roof forms a short-arm beam structure, and the deformation of the part of the roof is reduced. On the other hand, the broken swelling characteristic of rock is utilized to the maximum extent, the self bearing capacity of broken swelling waste rock is exerted, the sinking deformation of a high-level hard top plate is reduced, and the integral stability of the roadway is ensured. Therefore, reasonable kerf height is critical to the control of the portion of the top plate. The design height of the joint seam should satisfy the initial crushing and expansion balance equation of the collapsed waste rock, namely the volume of the roof rock mass after collapse and crushing and expansion in the height range of the joint seam is equal to the volume of the mined coal. The volume of the caving rock mass referred to herein should be the volume in the uncompacted state.
2) After the topping is carried out, the stress of the short-armed beam structure is actually optimized, and the deformation control of the short-armed beam structure is simplified. The most effective method is to use an anchor cable to bind the high-level stable rock stratum to ensure that the high-level stable rock stratum and the high-level stable rock stratum have synchronous deformation characteristics. As long as the anchor cable material is continuous, the short-arm beam structure can keep safe and stable. To achieve this, a constant-resistance anchor cable with high strength and capable of enduring large deformation without breaking is selected. Secondly, reasonable supporting strength is designed. Therefore, the support strength of the constant-resistance anchor cable can meet the static balance equation for keeping the short-arm beam structure stable.
3) In the dynamic pressure stage, the bearing structure movement cannot be completely prevented, and the sinking movement of the bearing structure inevitably causes the short-arm beam structure to be strongly pressed, so that the short-arm beam structure is easily broken. If there is no support in the roadway, the location of the short arm beam structure break must be above the roadway, which is dangerous. If the high-strength support is arranged in the roadway, the short-arm beam top plate is not broken or broken inside the solid coal side. Based on the thought, temporary high-strength dynamic pressure bearing support is timely carried out, and the key for ensuring the stability of the short-arm beam structure in the dynamic pressure period is provided. And the design of dynamic pressure bearing support at least meets the mechanical balance condition of the internal fracture of the coal side of the short-arm beam structure.
In view of the above problem, the embodiment of the present disclosure provides a control method for a coal mining roadway roof, as shown in fig. 1, the control method includes:
step S110, determining the joint cutting height according to the coal mining height, the roadway height and the initial crushing expansion coefficient of the roof rock mass;
and step S120, forming a top plate kerf on the roadway top plate according to the determined kerf height.
According to the control method of the coal mining roadway roof, the initial crushing expansion coefficient of the roof rock mass is considered in the design parameters of the actual upper seam height, and the changed crushing expansion coefficient of the gangue in the goaf gangue compaction process is not taken as the design basis. The change process of the residual crushing expansion coefficient and the crushing expansion coefficient influences the final deformation amount and the deformation process of the pressure-bearing structure, and the change process has no direct relation with the height of the cutting seam. Therefore, the height of the joint cutting is determined through the coal mining height, the roadway height and the initial crushing expansion coefficient of the top plate rock mass, the goaf can be filled completely just after the crushed expansion of the caving gangue, the top plate of the high-rise position is contacted with the gangue once sinking, and then the gangue is compacted continuously until the gangue is stable.
Specifically, as shown in fig. 2, the control method of the coal mining roadway roof further includes:
step S210, determining constant-resistance anchor cable supporting force of the top plate before the short-arm beam is broken according to anchor cable supporting force required before the arm beam is broken, roadway width, joint cutting height, joint cutting angle, distance between the action point of anchor cable supporting resultant force and roadway solid coal side, dead weight of the short-arm beam structure and supporting force of collapse waste rock on the short-arm beam structure in the vertical direction;
step S220, determining constant-resistance anchor cable supporting force after the short-arm beam of the top plate is broken according to anchor cable supporting force required after the arm beam is broken, the self weight of the short-arm beam structure, the distance between the position of a broken point of the short-arm beam structure and an entity coal wall, the position of the entity coal wall, supporting force of collapsed waste rocks on the short-arm beam structure in the vertical direction, the distance between the action point of the combined force of the anchor cable supporting and the entity coal wall of the roadway;
and step S230, according to the determined constant-resistance anchor cable supporting force required before and after the short-arm beam breaks, supporting the top plate within the range of the seam by using the constant-resistance anchor cable.
Specifically, as shown in fig. 3, the control method of the coal mining roadway roof further includes:
step S310, determining dynamic pressure bearing temporary supporting force according to dynamic pressure bearing temporary supporting force, roadway width, average volume weight of an overlying strata, volume weight of a short-arm beam structure, joint cutting height, joint cutting angle, supporting coefficient of caving gangue to a short-arm beam, stress concentration coefficient caused by mining, coal seam burial depth, number of each row of constant-resistance anchor cables, internal friction angle of the short-arm beam structure, cohesive force of the short-arm beam structure, pre-tightening force of a single constant-resistance anchor cable and supporting force of the caving gangue to the short-arm beam structure in the vertical direction;
and step S320, carrying out dynamic pressure bearing support on the top plate according to the determined dynamic pressure bearing temporary support force.
Hereinafter, each step of the control method of the coal mining roadway roof provided by the present disclosure will be described in detail.
In step S110, the kerf height is determined according to the coal mining height, the roadway height and the initial crushing expansion coefficient of the roof rock mass.
Specifically, the created truncated height calculation model is shown in fig. 7A and 7B. According to field test observation, the top plate in the cutting joint range can be preferentially collapsed in the initial stage of the movement of the top plate. The roof strata above the kerf has not yet moved significantly. If the height of the cutting seam is reasonably designed, the goaf can be completely filled right after the collapsed waste rock is crushed and expanded. Once the top plate of the high-rise position sinks, the top plate of the high-rise position is contacted with the gangue, and then the gangue is continuously compacted until the gangue is stabilized.
Therefore, the design parameters of the actual upper seam height mainly take the initial crushing expansion coefficient of the roof rock mass into consideration, and the crushing expansion coefficient of the gangue, which changes in the goaf gangue compaction process, is not taken as the design basis. The change process of the residual crushing expansion coefficient and the crushing expansion coefficient influences the final deformation amount and the deformation process of the pressure-bearing structure. This is not directly related to the kerf height. In addition, considering the situation that the mining height and the roadway height are not equal, the kerf height and the coal mining height should satisfy the following relation:
wherein HfIs the height of the cutting seam HmFor coal mining height, HrIs the height of the roadway, K0Is the initial crushing and swelling coefficient of the roof rock mass.
Under the action of the pressure of the overburden rock layer, the volume of the collapsed gangue is further compressed. In the whole process, the surrounding rock of the roadway can gradually deform under the sinking and extruding action of the rock block B. However, the rock block B is always in a gangue supporting state, so that the sudden mine pressure display which is too violent cannot be generated in the roadway. After the waste rock in the goaf is compacted, the rock block B forms a stable structure under the combined action of the coal body, the waste rock in the goaf and the adjacent rock blocks A and C.
In step S120, a roof kerf is formed on the roadway roof according to the determined kerf height.
Specifically, the kerf height H is determinedfAnd then, forming a top plate kerf on the roadway top plate according to the determined kerf height. The particular top panel slitting method of implementation is conventional in the art and will not be described in detail herein.
In step S210, the constant-resistance anchor cable supporting force of the top plate before the short-arm beam is broken is determined according to the anchor cable supporting force required before the arm beam is broken, the roadway width, the joint cutting height, the joint cutting angle, the distance between the acting point of the anchor cable supporting resultant force and the roadway solid coal wall, the self weight of the short-arm beam structure, and the supporting force of the collapse gangue on the short-arm beam structure in the vertical direction.
Specifically, according to the stable characteristic of the pillar-free self-entry roof structure, the high-level broken rock beam above the joint seam can keep self-stability under the combined action of the support force of the collapsed waste rock and the support force of the coal body. Therefore, for the short-arm beam structure below the cutting seam, the short-arm beam structure can be anchored into the stable rock stratum above the short-arm beam structure by adopting the anchor cable with high strength and high prestress, so that the short-arm beam structure and the stable rock stratum have synchronous deformation characteristics, and the stability of the short-arm beam structure can be realized. The supporting material is required to be capable of applying high pretightening force, so that the self-bearing characteristic of the surrounding rock is fully exerted, and the integral stability of the surrounding rock is ensured.
In addition, certain friction effect can be produced to the tunnel roof usually in the goaf waste rock caving process, leads to the tunnel roof to easily present big deformation characteristic. This requires that the roof support material must also have impact-resistant, stretch-resistant characteristics to ensure that the support material does not break when large deformations occur. Therefore, the common anchor rod can not meet the characteristics, and the top plate is reinforced by adopting a full-section high-prestress constant-resistance large-deformation anchor cable supporting technology. Compared with the common anchor cable, the constant-resistance large-deformation anchor cable has the main advantages of high prestress, high elongation, impact resistance, capability of absorbing deformation energy of surrounding rocks and the like, and the working principle is shown in fig. 7A and 7B.
The constant-resistance anchor cable parameter design calculation model is shown in fig. 8A and 8B. In the figure, G is the self gravity of the short-arm beam structure; fcAnchor cable support force. It should be noted that the positions of the force and resultant action points of the anchor line are simplified since the balance of the whole short-arm beam is considered here. L iscThe distance between the acting point of the combined force of the anchor cable support and the entity coal side of the roadway, wherein a is the width of the roadway and HfAnd beta is the kerf height and angle, x, respectively0The distance between the position of the short-arm beam structure fracture point and the solid coal side, qcAnd σ0Respectively the bearing pressure, K, of the solid coal side position and the short-arm beam breaking point position1The support coefficient (capable of taking 0-1) of the collapsed waste rock to the short-arm beam is gamma1The average volume weight of the overlying rock stratum is shown, and H is the buried depth of the coal bed.
Supporting force F required before fracture of short-arm beam structurec1：
According to fig. 8A, the following static equilibrium equation should be satisfied before the short-arm beam structure breaks:
wherein, F0The supporting force of the solid coal side position is changed along with different anchor cable supporting strengths; fgIn order to support the short-arm beam structure by the collapsed gangue in the vertical direction, the support force can be expressed by the following formula:
Fg＝K1γ1HHftanβ (3)
according to the calculation formula, the top plate in the joint cutting range can be kept stable when the design strength of the constant-resistance anchor cable meets the following conditions:
according to the formula (4), the optimal combined relation of the anchor cable support resistance and the row spacing can be calculated and obtained. It should be noted that, theoretically, when the anchor cable is installed at the end position of the short-arm beam, the required supporting resistance is the minimum. However, in practical engineering, while saving the resistance of the anchor cable support, the spacing between the anchor cables, the support area, and the coordination between the anchor cable support and the coal mining process should be considered. Therefore, the anchor cable support resistance and the spacing need to be considered in a coordinated manner rather than calculated independently.
In step S220, the constant-resistance anchor cable supporting force after the short-arm beam of the top plate is broken is determined according to the anchor cable supporting force required after the arm beam is broken, the self weight of the short-arm beam structure, the distance between the position of the broken point of the short-arm beam structure and the solid coal wall, the position of the solid coal wall, the supporting force of the collapsed gangue on the short-arm beam structure in the vertical direction, the distance between the acting point of the anchor cable supporting resultant force and the solid coal wall of the roadway.
Specifically, the supporting force F required after the short-armed beam structure is brokenc2：
According to fig. 8A, the following static equilibrium equation should be satisfied after the short-arm beam structure is broken:
solving the equation set (5) to obtain that the support strength of the anchor cable meets the following conditions:
in step S230, a constant-resistance anchor cable is used to support the top plate within the range of the seam according to the determined constant-resistance anchor cable supporting force required before and after the short-arm beam breaks.
Specifically, according to equations (4) and (6), it can be obtained that the anchor cable support resistance which meets the stability condition before and after the short-arm beam breaks at least should satisfy:
in step S310, the dynamic pressure bearing temporary supporting force is determined according to the dynamic pressure bearing temporary supporting force, the roadway width, the average bulk density of the overlying strata, the bulk density of the short-arm beam structure, the joint cutting height, the joint cutting angle, the support coefficient of the collapsed gangue to the short-arm beam, the stress concentration coefficient caused by mining, the coal seam burial depth, the number of each row of constant-resistance anchor cables, the internal friction angle of the short-arm beam structure, the cohesion of the short-arm beam structure, the pre-tightening force of a single constant-resistance anchor cable, and the support force of the collapsed gangue to the short-arm beam structure in the vertical direction.
The roadway formed by the pillar-free self-roadway mining is positioned at the edge of the goaf. Under the influence of the rotary motion of a high-level hard top plate, the roadway at the initial stage of roadway forming usually shows a relatively obvious dynamic pressure phenomenon, and the top plate can generate relatively obvious deformation. If the deformation of the top plate can not be effectively controlled in the period, the rock mass of the top plate is damaged, and the stability of the roadway is not facilitated. Therefore, when the non-coal-pillar self-entry mining technology is adopted, on the basis of constant-resistance large-deformation anchor cable support, a dynamic pressure bearing support structure (as shown in fig. 9) is adopted to temporarily support the roadway during the dynamic pressure influence period so as to prevent the low-level roof of the roadway from being severely deformed or even unstable due to the influence of the movement of the high-level roof. If support the reasonable of intensity design, can guarantee that the fracture position of short armed roof beam roof is located the inboard of coal group, this is very crucial to the stability of later stage roof. Therefore, the design of the dynamic pressure bearing support parameters is based on the mechanical balance condition that the short arm beam is broken at the inner side of the coal upper.
Based on the above discussion, assuming that the fracture position of the short-arm beam top plate is located right above the solid coal side (i.e. the critical value of the required dynamic pressure bearing temporary support force), the dynamic pressure bearing support parameters in this state can be obtained by using the upper limit analysis theory. In order to obtain a design method of dynamic pressure bearing support, the surrounding rock of the top plate of the short-arm beam is regarded as a homogeneous rock mass, and a theoretical calculation model is established according to the fracture and damage characteristics of the short-arm beam, as shown in fig. 9. In the figure, f (x) is a fracture curve equation of the short-arm beam structure; the shear stress at the fracture surface of the short-armed beam is taunPositive stress of σnAnd run alongCurve f (x) is uniformly distributed.
Assuming that the damage of the short-arm beam structure rock mass meets the Hoek-Brown strength criterion:
τn＝Aσc[(σn+σt)σc -1]B (8)
based on the Hoek-Brown strength criterion and the limit analysis upper limit method in plastic mechanics, the relationship between the fracture position of the short-arm beam structure and several key parameters is obtained by solving in the previous research:
in the formula, L is the horizontal distance between the fracture position of the short-arm beam structure and the solid coal side, and gamma is2Volume weight, K, for short-armed beam structures2N is the number of the anchor cables in each row, f is the stress concentration coefficient caused by miningpIs the pretightening force of a single constant-resistance anchor cable, d is the row spacing of the anchor cables, c is the cohesive force of a short-arm beam structure,
internal angle of friction, F, for short-armed beam constructiontAnd bearing temporary supporting force for dynamic pressure.
In order to obtain the critical value of the dynamic pressure bearing support force required when the fracture position of the short-arm beam structure is located right above the solid coal upper, it is only required that L in the formula (9) is equal to zero. The calculation method for obtaining the dynamic pressure bearing temporary support force based on the method comprises the following steps:
in step S320, the top plate is subjected to dynamic pressure bearing support according to the determined dynamic pressure bearing temporary support force.
Specifically, dynamic pressure bearing support is carried out on the top plate according to the determined dynamic pressure bearing temporary support force. The concrete dynamic pressure bearing support method is a conventional means in the field and will not be described in detail.
By combining the research and calculation results, the comprehensive design method for controlling the gob-side entry retaining roof by the N00 construction method can be obtained. The method mainly comprises three aspects, namely directional roof cutting height design, constant-resistance anchor cable support strength design and dynamic pressure bearing support strength design. The three should satisfy the following conditions simultaneously:
in the practical application process, the roof joint-cutting design needs to be based on the premise of ensuring the joint-cutting effect. If the joint cutting effect can not meet the design requirement, the joint cutting height should be properly increased to increase the rock mass caving height.
In addition, for the sake of safety, the required constant-resistance anchor cable supporting force before and after the determined short-arm beam is broken is multiplied by the safety factor to form the required constant-resistance anchor cable supporting force; and multiplying the determined dynamic pressure bearing temporary support force by the safety factor to form the required dynamic pressure bearing temporary support force. Namely, the design of constant-resistance anchor cable support and dynamic pressure bearing support should consider a certain safety factor, and the value of the safety factor is 1.5-2.
Next, a control method of the coal mine tunnel roof will be exemplarily described.
Engineering background
In an example, the test working face of the industrial test is an S1201-II working face, the inclined length of the test working face is 280m, the strike length is 2344m, the thickness of a coal seam is 3.81-4.35m, the average thickness of the coal seam is 4.11m, the burial depth is 115-170 m, the occurrence of the coal seam is stable, and the inclination angle of the coal seam is nearly horizontal. The face lane layout is shown in fig. 10. In the figure, an S1201-II glue transportation gateway (tunnel driving) is a tunnel which is tunneled in advance, and a return air gateway (tunnel retaining) is a tunnel which is formed by mining coal and forming the same side by adopting a roof cutting pressure relief non-coal-pillar self-roadway mining technology in the working face stoping process, and the total length is 2344 m.
Kerf parameter design
According to the rock mass crushing and swelling characteristics, the volume of the rock mass after the rock mass strides can be increased. If the caving rock mass can be filled with the mining space, the caving waste rock can effectively support the overlying rock stratum, so that the caving waste rock is controlled to sink. Therefore, the height of the top plate joint cutting is not less than the height of the full mining space after the top plate joint cutting is collapsed theoretically, calculation can be carried out through a formula (1), and parameter values are shown in a table 1.
Table 1 formula (1) parameter values:
through calculation, the optimal kerf height H can be obtainedf8.72 m. Therefore, the slit depth is designed to be 9 m. In order to reduce the friction effect in the process of gangue caving, the cutting joint angle is designed to be 10 degrees.
Constant-resistance anchor cable support parameter design
The values of the parameters of the formula (7) are shown in table 2 according to the engineering geological conditions of the test working face.
TABLE 2 value of formula (7) parameter
Among them, it should be noted that K is the most dangerous state in consideration of the case where the slitting effect is not good1Take 0. In addition, for the sake of safety, the supporting force at the solid coal upper position is regarded as 0. According to the formula (7), when LcWhen different values are taken, F can be obtained respectivelyc1And Fc2As shown in fig. 11.
It can be seen from the figure that the closer the anchor cable support resultant force position is to the goaf side, the smaller the required support force is. The combination mode of the anchor cable supporting resultant force meeting the design requirements and the distance between the anchor cable supporting resultant force and the solid coal roadway side has multiple choices. In fact, as long as the reasonable size requirement of support is met, the combination modes of the diameter of the anchor cable, the row spacing among the anchor cables and the like are also possible.
The arrangement mode of the anchor cable is selected by considering various factors such as supporting area, construction process and the like. In the test, the overall equipment layout is considered, and 5 constant-resistance anchor cables are adopted for supporting the roadway in each section design. The row pitch of the anchor cables is 0.8m (coal mining step pitch), the diameter of the anchor cables is 21.8mm, and the breaking force is 530 kN. The length of the anchor cable is generally equal to b + (1.0-1.5) m, and the test is 10.5 m. The space between the anchor cables is respectively designed to be 1245m, 1295m, 1230m and 1230m according to the arrangement space of the hydraulic support and the drilling machine.
Namely, the distance between the resultant equivalent action point of the anchor cable support and the goaf is 2500 mm. Substituting the parameters into the formula (7) can obtain the required anchor cable supporting force of the roadway with unit length of 2158.3 kN. The constant-resistance anchor cable supporting force adopted by the embodiment is sigma fciAnd/0.8 m is 3312.5 kN. The safety coefficient is 1.5, and the safety production requirement can be met. The design parameters are shown in fig. 12.
Dynamic pressure bearing support parameter summation
The values of the parameters of equation (10) are shown in table 3, based on the geological report. .
TABLE 3 values of the parameters of equation (10)
Based on the parameters, the dynamic pressure bearing temporary support force required by the unit length roadway is calculated to be 2191.9kN, and the dynamic pressure coefficient is considered to be 2.0.
In this embodiment, two rows of dynamic pressure bearing supports are designed for each section. The working resistance of the dynamic pressure bearing support is designed to be 4000kN, the support area is 1950mm (L) multiplied by 1200mm (W), and the distance between the two brackets is 2.4m, as shown in figure 13. The supporting resistance of the roadway in unit length is 3333.3kN, the safety coefficient is 1.5, and the requirement of safe production is met. According to the mine pressure monitoring result, the top plate is gradually stabilized after 160m behind the working face, and then the pressure bearing support can be gradually withdrawn.
In the application process of the coal mine, the comprehensive design method for controlling the top plate of the N00 construction method is used for monitoring the crack rate in a cutting and drilling hole, the stress of a constant-resistance anchor cable and the stress of a dynamic pressure bearing support respectively, as shown in fig. 14 (the ordinate is the accumulative approach of the top and bottom plates (Roof-to-floor deformation), and the abscissa is the mining distance of a working face (Position of gob-side entry)).
According to the monitoring data, the maximum value of the accumulated approaching amount of the top and bottom plates is 149mm within the range of 0-900 m of the entry retaining, the average value is 101mm, and the deformation amount is generally small. In addition, the average crack rate in the directional roof cutting hole is more than 80%. The stress of the constant-resistance anchor cable is stabilized at about 300 kN. The maximum stress of the dynamic pressure bearing support is 40MPa, and the dynamic pressure bearing support enters a stable state about 160m behind the working surface. After the comprehensive design method and design parameters for top plate control are adopted, the deformation control effect of the top plate is good, and the requirement of safe production is completely met.
It should be noted that although the steps of the control method of the coal mine roof in the present disclosure are described in the figures in a particular order, this does not require or imply that the steps must be performed in that particular order or that all of the illustrated steps must be performed to achieve the desired results. Additionally or alternatively, certain steps may be omitted, multiple steps combined into one step execution, and/or one step broken down into multiple step executions, etc.
Other embodiments of the disclosure will be apparent to those skilled in the art from consideration of the specification and practice of the disclosure disclosed herein. This application is intended to cover any variations, uses, or adaptations of the disclosure following, in general, the principles of the disclosure and including such departures from the present disclosure as come within known or customary practice within the art to which the disclosure pertains. It is intended that the specification and examples be considered as exemplary only, with a true scope and spirit of the disclosure being indicated by the following claims.
It will be understood that the present disclosure is not limited to the precise arrangements that have been described above and illustrated in the drawings, and that various modifications and changes may be made without departing from the scope thereof, which is limited only by the claims appended hereto.
Claims (9)
1. A control method for a coal mining roadway roof is characterized by comprising the following steps:
determining the height of a joint according to the coal mining height, the roadway height and the initial crushing expansion coefficient of a roof rock mass;
forming a top plate kerf on the top plate of the roadway according to the determined kerf height;
determining constant-resistance anchor cable supporting force of the top plate before the short-arm beam is broken according to anchor cable supporting force required before the short-arm beam is broken, roadway width, joint cutting height, joint cutting angle, distance between acting point of anchor cable supporting resultant force and roadway solid coal side, dead weight of the short-arm beam structure, and supporting force of collapsed waste rock on the short-arm beam structure in the vertical direction;
determining constant-resistance anchor cable supporting force of the broken short-arm beam of the top plate according to anchor cable supporting force required after the short-arm beam is broken, the self weight of the short-arm beam structure, the distance between the position of a broken point of the short-arm beam structure and an entity coal side, the position of the entity coal side, supporting force of collapsed waste rock on the short-arm beam structure in the vertical direction, and the distance between the action point of the combined anchor cable supporting force and the entity coal side of the roadway;
and according to the determined constant-resistance anchor cable supporting force required before and after the short-arm beam is broken, supporting the top plate within the range of the seam by using the constant-resistance anchor cable.
2. The control method as claimed in claim 1, wherein the kerf height is:
wherein HfIs the height of the cutting seam HmFor coal mining height, HrIs the height of the roadway, K0Is the initial crushing and swelling coefficient of the roof rock mass.
3. The control method according to claim 1, wherein the constant-resistance anchor cable support force required before the short-arm beam breaks is:
wherein, Fc1Required before the short-arm beam is brokenConstant-resistance anchor cable supporting force, a is the width of the roadway and HfIs the height of the cutting seam, beta is the angle of the cutting seam, LcThe distance between the acting point of the resultant force of the anchor cable support and the solid coal wall of the roadway, G is the self weight of the short-arm beam structure, and F is the distance between the acting point of the resultant force of the anchor cable support and the solid coal wall of the roadwaygThe support force of the short-arm beam structure in the vertical direction is realized for the collapse waste rock.
4. The control method according to claim 1, wherein the constant-resistance anchor cable support force required after the short-arm beam is broken is:
wherein, Fc2The supporting force of the constant-resistance anchor cable required after the short-arm beam is broken is G, the self weight of the short-arm beam structure is X0The distance between the position of the short-arm beam structure fracture point and the solid coal side, qcFor supporting pressure at the position of the solid coal side, FgSupporting force of the short-armed beam structure in the vertical direction for caving waste rock, LcThe distance between the acting point of the combined force of the anchor cable support and the entity coal side of the roadway, wherein a is the width of the roadway and HfIs the kerf height and beta is the kerf angle.
5. The control method according to claim 1, characterized by further comprising:
determining dynamic pressure bearing temporary supporting force according to dynamic pressure bearing temporary supporting force, roadway width, average volume weight of overlying strata, volume weight of a short-arm beam structure, joint cutting height, joint cutting angle, support coefficient of caving gangue to a short-arm beam, stress concentration coefficient caused by mining, coal seam burial depth, number of each row of constant-resistance anchor cables, internal friction angle of the short-arm beam structure, cohesion of the short-arm beam structure, pretightening force of a single constant-resistance anchor cable and supporting force of the caving gangue to the short-arm beam structure in the vertical direction;
and carrying out dynamic pressure bearing support on the top plate according to the determined dynamic pressure bearing temporary support force.
6. The control method according to claim 5,
wherein, FtFor dynamic pressure bearing temporary support force, a is roadway width, gamma1Is the average volume weight of overburden, gamma2Is the volume weight of the short-armed beam structure, HfIs the height of the cutting seam, beta is the angle of the cutting seam, K1Is the support coefficient of the collapsed waste rock to the short arm beam, K2H is the coal seam burial depth, n is the number of each row of constant-resistance anchor cables,
is the internal friction angle of the short-armed beam structure, c is the cohesion of the short-armed beam structure, fpPre-tightening force for single constant-resistance anchor cable, FgAnd d is an anchor cable row pitch for supporting the short-arm beam structure in the vertical direction by the collapsed waste rock.
7. The control method according to claim 1, wherein the determined constant-resistance anchor rope supporting force required before and after the short-arm beam is broken is multiplied by a safety factor to form the required constant-resistance anchor rope supporting force.
8. The control method according to claim 5, wherein the determined dynamic pressure bearing temporary support force is multiplied by a safety factor to form a required dynamic pressure bearing temporary support force.
9. The control method according to claim 7 or 8, characterized in that the safety factor is 1.5-2.
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| CN113669063A (en) * | 2021-10-22 | 2021-11-19 | 中国矿业大学(北京) | Roof cutting self-entry surrounding rock control method |
| CN113914861A (en) * | 2021-12-15 | 2022-01-11 | 中国矿业大学(北京) | Non-pillar self-entry balanced mining method and equipment |
| CN114419982B (en) * | 2021-12-29 | 2024-03-15 | 山东科技大学 | Model test system and method for deformation and damage of goaf roadway of coal pillar reserved in soft rock stratum |
| CN115467662B (en) * | 2022-11-15 | 2023-01-24 | 中国矿业大学(北京) | Rock mass crushing and swelling characteristic in-situ test and evaluation method |
| CN115659706B (en) * | 2022-12-28 | 2023-04-04 | 北京科技大学 | Design method for roof joint cutting in coal pillar-free self-entry mining |
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