AU2021100425A4 - Method for Quantitatively Designing Roadway Support Based on Size of Plastic Zone of Surrounding Rock - Google Patents
Method for Quantitatively Designing Roadway Support Based on Size of Plastic Zone of Surrounding Rock Download PDFInfo
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- 230000015572 biosynthetic process Effects 0.000 claims description 40
- 238000005755 formation reaction Methods 0.000 claims description 40
- 239000000725 suspension Substances 0.000 claims description 12
- 238000006073 displacement reaction Methods 0.000 claims description 9
- 238000007586 pull-out test Methods 0.000 claims description 7
- 230000002457 bidirectional effect Effects 0.000 claims description 6
- 238000005507 spraying Methods 0.000 claims description 6
- 230000008961 swelling Effects 0.000 claims description 6
- 230000001131 transforming effect Effects 0.000 claims description 6
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- 239000003245 coal Substances 0.000 description 6
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- 238000004458 analytical method Methods 0.000 description 3
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- 238000009826 distribution Methods 0.000 description 3
- 238000005065 mining Methods 0.000 description 3
- 238000013528 artificial neural network Methods 0.000 description 2
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- 230000006641 stabilisation Effects 0.000 description 2
- 238000011105 stabilization Methods 0.000 description 2
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Classifications
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- E—FIXED CONSTRUCTIONS
- E21—EARTH OR ROCK DRILLING; MINING
- E21D—SHAFTS; TUNNELS; GALLERIES; LARGE UNDERGROUND CHAMBERS
- E21D9/00—Tunnels or galleries, with or without linings; Methods or apparatus for making thereof; Layout of tunnels or galleries
- E21D9/14—Layout of tunnels or galleries; Constructional features of tunnels or galleries, not otherwise provided for, e.g. portals, day-light attenuation at tunnel openings
-
- E—FIXED CONSTRUCTIONS
- E21—EARTH OR ROCK DRILLING; MINING
- E21C—MINING OR QUARRYING
- E21C41/00—Methods of underground or surface mining; Layouts therefor
- E21C41/16—Methods of underground mining; Layouts therefor
- E21C41/18—Methods of underground mining; Layouts therefor for brown or hard coal
-
- E—FIXED CONSTRUCTIONS
- E21—EARTH OR ROCK DRILLING; MINING
- E21F—SAFETY DEVICES, TRANSPORT, FILLING-UP, RESCUE, VENTILATION, OR DRAINING IN OR OF MINES OR TUNNELS
- E21F17/00—Methods or devices for use in mines or tunnels, not covered elsewhere
-
- E—FIXED CONSTRUCTIONS
- E21—EARTH OR ROCK DRILLING; MINING
- E21D—SHAFTS; TUNNELS; GALLERIES; LARGE UNDERGROUND CHAMBERS
- E21D21/00—Anchoring-bolts for roof, floor in galleries or longwall working, or shaft-lining protection
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- Engineering & Computer Science (AREA)
- Mining & Mineral Resources (AREA)
- Life Sciences & Earth Sciences (AREA)
- General Life Sciences & Earth Sciences (AREA)
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- Devices Affording Protection Of Roads Or Walls For Sound Insulation (AREA)
Abstract
The present invention discloses a method for quantitatively designing a roadway support based on
a size of a plastic zone of a surrounding rock. The method includes the following steps of: step 1:
carrying out calculation to determine a size of a radius R of a plastic zone of a surrounding rock of a
roadway; step 2: classifying the surrounding rock of the roadway according to a calculation result; step
3: carrying out targeted support parameter design on different types of surrounding rocks according to
a classification result of the surrounding rock of the roadway; and step 4: modifying support parameters
according to on-site monitoring feedback information. According to the method, the support parameters
designed on the basis of the size of the plastic zone are scientific and reasonable, and classification of
the surrounding rock of the roadway is carried out by taking a radius size of the plastic zone as a main
index on the basis of obtaining the radius of the plastic zone of the roadway and geometric
characteristics of a boundary, so that targeted support parameter design is carried out on different types
of surrounding rocks so as to ensure the scientificity and economy of roadway support design. For
roadway support, the method is capable of playing roles in optimizing the support parameters of the
surrounding rock and preventing the roadway from roof fall accidents.
3/3
Anchorage cable
p21.6x8250mm
Rebar anchorage rod p20x2400mm
(an anchorage rod and a roof
anchorage rod cable are arranged in
:rows)
FIG. 4
Anchorage rope
p21.6x8250mm
Rebar anchorage rod (p20x2400mm
(an anchorage rod of a ledge and a
roof anchorage rod cable are
arranged in rows)
FIG. 5
Description
3/3
Anchorage cable p21.6x8250mm Rebar anchorage rod p20x2400mm (an anchorage rod and a roof anchorage rod cable are arranged in :rows)
FIG. 4
Anchorage rope p21.6x8250mm Rebar anchorage rod (p20x2400mm (an anchorage rod of a ledge and a roof anchorage rod cable are arranged in rows)
FIG. 5
Method for Quantitatively Designing Roadway Support Based on Size of Plastic
Zone of Surrounding Rock
[01] The present invention relates to the field of supporting coal mine roadways, and in particular, to a method for quantitatively designing a roadway support based on a size of a plastic zone of a surrounding rock
[02] China introduced roadway anchorage rod support technology in some state owned mines in the 1950s. Over the past decades, with the continuous development and improvement of anchorage rod support theory and anchorage rod support materials, anchorage rod support has become mainstream technology of coal mine roadway support. However, the design of anchorage rod support in China only stays at design methods such as engineering analogy, resulting in poor safety and reliability of a roadway, and hidden dangers that threaten the normal production of coal mines, which greatly restricts the development of anchorage rod support technology. Support design methods need to be improved, therefore, the development of a more scientific, practical and reliable anchorage rod support design method is imminent. At present, many scholars at home and abroad have conducted in-depth research on roadway support design methods and have obtained many useful research results.
[03] A method for dynamically designing a coal road anchorage rod support including "crustal stress testing-geomechanics evaluation-initial design-field monitoring-feedback information-modification design" is proposed. According to a relationship between a horizontal principal stress and a vertical stress, a formation stress state is divided into four situations. Based on this, a "rigid" beam theory based on high level crustal stress and a "rigid" wall theory based on a vertical crustal stress are proposed, and the corresponding method for designing an anchorage rod parameter is established.
[04] It is believed that the damage of a rock mass medium of a deep soft rock roadway engineering is mainly caused by an uncoupling between a support body and a surrounding rock. Based on this, a method for non-linearly designing anchorage-net cable coupling support that can not only fully utilize an ability of an anchorage net to actively support a shallow surrounding rock can be but also mobilize a support capacity of a deep surrounding rock by an anchorage cable is proposed. Based on analysis of an anchorage mechanism, the present invention proposes a theory of high prestress and strong support, emphasizes the decisive role of an anchorage rod prestress and the diffusion of the prestress, uses a method of combining fuzzy mathematics and hierarchy analysis, considers the existing experience of a bolt-shotcrete support, proposes a fuzzy empirical analysis method for an analog design of the bolt-shotcrete support, regards a thickness value of a loose circle as a manifestation of a multi-factor comprehensive index, and combines with a loose circle theory, and uses the thickness value of the loose circle to classify the surrounding rock of the loose circle. Based on this, a method for designing a support is proposed and a corresponding support design software is developed. Based on the principle of an artificial neural network, a multi-level artificial neural network for anchorage rod support design is constructed. Based on the classification of the surrounding rock, a coal road anchorage rod support form and a parameter optimization design are constructed. Based on a principle of a self stabilization invisible arch and on a basis of summarizing a mechanism of reinforcement of the surrounding rock by an anchorage rod, a concept that a limit self stabilization invisible arch is the maximum interval boundary where the surrounding rock of an underground roadway may be unstable is proposed, and a corresponding anchorage rod support design method is established. In summary, the forgoing studies have carried out an in-depth study of a roadway support design method from the perspectives of a support design theory, a support parameter design method, and the relationship between the surrounding rock and the support body.
[05] However, most of the surrounding rocks of underground coal mine roadways are coal-measure sedimentary rocks. Because of the complexity and variability of sedimentary rocks, coupled with an influence of a superimposed stress field of an original crustal stress field and a mining stress field generated by mining, when the existing support method performs support parameter design, only one support parameter is adopted for one roadway, making it unable to adapt to the ever-changing support requirements of the surrounding rock, resulting in excess or insufficient support strength of the roadway, which makes protection economy or safety of a roadway support become worse. Therefore, the present invention proposes a method for quantitatively designing a roadway support based on a size of a plastic zone of a surrounding rock, which can realize the precise and targeted support quantitative design of the roadway.
[06] The objective of the present invention is to overcome the shortcomings of the prior art and provide a method for quantitatively designing a roadway support based on a size of a plastic zone of a surrounding rock to solve the above technical problems.
[07] In order to achieve the forgoing objectives, the present invention provides the following solutions:
[08] The method for quantitatively designing the roadway support based on the size of the plastic zone of the surrounding rock includes the following steps:
[09] step 1: carrying out calculation to determine a size of a radius R of a plastic zone of a surrounding rock of a roadway;
[010] step 2: classifying the surrounding rock of the roadway according to a calculation result;
[011] step 3: carrying out targeted support parameter design on different types of surrounding rocks according to a classification result of the surrounding rock of the roadway; and
[012] step 4: modifying support parameters according to on-site monitoring feedback information.
[013] As a further solution of the present invention, a method for calculating the radius R of the plastic zone of the surrounding rock in the step 1 is as follows: for a circular roadway, relative parameters are substituted into the following formula (12) and formula (13) to directly calculate the size of the radius R of the plastic zone. When a roadway section is rectangular, arched or other shapes, FLAC3D, UDEC and other computer numerical simulation software are used to obtain the radius R of the plastic zone ;
[014] A stress solution of a circular roadway of the surrounding rock under a bidirectional unequal pressure stress field is as follows;
=- r(+) 1+r -I(A- ) 1-4a+3- cos 20
r 0= (-1) 1+2 -3 sin 20
[015] (1)
[016] where, 7 is a radial stress at any point; is a hoop stress at any point;
rO is a shear stress at any point; y is a bulk density of the rock; H is a buried depth
of the roadway; X is a lateral pressure coefficient; a is radii r of a circular roadway, a polar coordinate of any point;
[017] In order to obtain a principal stress at a certain point in the surrounding rock of the circular roadway, a stress formula under a polar coordinate system needs to be transformed into a stress formula under a rectangular coordinate system using the following formula:
UrCI+ 2 r cos 2O-r, sin 20 2 U = ' 9 - ' 0cos20+r sin2O 2 2 r- ca sin20+r cos20 2
[018] (2)
[019] A calculation formula of the principal stress in elasticity is as follows:
I] Ox +4y ++4
2 2 "3 _ 2 /(7 2 + 4
[020] (3)
[021] After substituting formula (1) into formula (2) and then substituting formula (1) into formula (3), the principal stress at any point in the surrounding rock of the circular roadway expressed in polar coordinates is as follows:
S = 2TO + 1 (, -r,)2 +4r 2 2 2
, C3 + 2 2 2 (4)
[022]
[023] A stress state of one point has three main directions. When the stress state of one point is expressed by the principal stress, the stress can be expressed as:
7 ( i:)m + i
[024] (5)
[025] where Si is a main deviation stress, and am-is an average stress;
C7+ 23
[026] (6)
[027] When studying the stress of the surrounding rock of the roadway, the stress is simplified as a plane problem. Therefore, according to formula (5) and formula (6), a calculation formula of a principal deviation stress is obtained:
= r ]2u 3- u3
= 2c 3 -n
[028] (7)
[029] Formula (4) is substituted into formula (7) to obtain:
7r+ 7i ++ ,rt 6 2
6 2
[030]
S +S3 -U Gr ± JT
[031] 3 (9)
[032] According to a ultimate principal stress at any point and combined with Mohr-Coulomb strength criterion, the Coulomb-Mohr strength criterion expressed by the ultimate principal stress ai and (3 is obtained, which is ultimate equilibrium condition:
cr, o=2Ccos( =1 2C co o +1+sing~a + --+in , e
[033] In~ p 1Sino (10)
[034] Where C is a cohesive force of an elastoplastic medium; <p is an internal friction angle of the elastoplastic medium; after transforming the formula (10), the following is obtained:
[035] (2a, - cj)+( 2 cr - a)-(at + q3 )sin ip - 2C Cos(p
[036] After further transforming the above formula, the following is obtained:
[037] 31t + 3s2 -(a +or,) sin (o=2CCOSp
aIL 2 7 H(1-A)sinqpcos2- =2Ccosqp+7 H(1+2)sing-3(s,+s,)
[038] r (11)
[039] Therefore, based on formula (11), the following is obtained:
r - 2a 2 7H sin Tcos 20(1-2)
[040] 2Ccosf+yH(l+A)sinip-3(s, +s) (12)
[041] Formula (12) is the solution of a radius of the minimum plastic zone of the surrounding rock of the circular roadway when the lateral pressure coefficient ) is not equal to 1;
[042] When the lateral pressure coefficient k=1, the formula (11) becomes:
[043] 2Ccos(±2YHsin(-3(s 1 ±S3 )=0
[044] Formula (9) is substituted into the forgoing formula, the following is obtained:
7H r =a
[045] 2(C +2)H sin p 13)
[046] Formula (13) is the solution of the radius of the plastic zone of the surrounding rock of the circular roadway when the lateral pressure coefficient is equal to 1.
[047] As a further solution of the present invention, in step 2, the method for classifying the surrounding rock of the roadway is as follows:
[048] After obtaining the radius R of the plastic zone of the surrounding rock, using the following formula to calculate the maximum hidden danger depth of a rock formation of a roof of the roadway:
H = Max(R e sin 0) -a ( 14)
[049] OCtO0r]
[050] where, 0 is an angle between the radius of the largest plastic zone of the surrounding rock and a horizontal direction; and a is a vertical distance between a center of the roadway and a surface of the roof;
[051] After obtaining the maximum roof fall hidden hazard depth H, the roof is divided into four categories according to a size of H:
[052] Class I: a stable roof H<1.5m, this kind of rock formation of the roof has the smallest plastic zone range, and the maximum roof fall hidden hazard height does not exceed 1.5m. A main control object of the support design is the broken dangerous rock block;
[053] Class II: a moderately stable roof 1.5m < H < 3.5m. This type of the rock
formation of the roof has the larger plastic zone range, and the maximum roof fall hidden hazard height does not exceed 3.5m. The roof produces a certain amount of deformation, and a main damage mode of the rock formation of the roof is a beam-type span, so the main control object of a support design is an unstable rock formation with the beam-type span on the roof;
[054] Class III: an unstable roof 3.5m < H < 5.5m. This type of the rock formation
of the rock has the larger plastic zone range, and the maximum roof fall hidden hazard height does not exceed 5.5m. The roof usually undergoes a large deformation, resulting in a certain amount of roof collapse. a main object of the support design is a cracked rock that is easy to fall;
[055] Class IV: an extremely unstable roof H>5.5m. This type of the rock formation of the roof has the largest plastic zone, and the maximum roof fall hidden hazard height exceeds 5.5m. This is mainly due to the large nonlinear deformation of the surrounding rock, which makes the roof prone to arched caving. So the main control object of the support design is a broken rock under a vault;
[056] Similarly, the maximum radius of the plastic zone of the ledge part of the surrounding rock of the roadway is obtained according to the following formula:
W= Max (R e sin ) -w Og_ -T_ Z. (15)
[057]
[058] where: w-a half width of the roadway;
[059] After obtaining the maximum radius W of the plastic zone of the ledge part, combined with a support theory of the loose circle of the surrounding rock in the prior art, the surrounding rock of the ledge part of the roadway is divided into four categories:
[060] Class I: a stable surrounding rock W<0.4m. The surrounding rock has a small swelling deformation. Generally, this kind of surrounding rock does not need anchorage rod support, and is supported by auxiliary support structures such as naked or steel mesh;
[061] Class II: a moderately stable surrounding rock 0.4m < W < 1.5m. this kind
of surrounding rock has a large amount of fracture and deformation, resulting in poor stability on both sides of the roadway. Therefore, a conventional anchorage rod is used for support;
[062] Class III: a unstable roof 1.5m < W < 3m. this kind of surrounding rock has
a wide crushing range, a large crushing depth, and a large deformation of the surrounding rock, resulting in obvious appearance of a bulge of the ledge part, poor stability of the two ledges, and anchorage net spraying is required for support and protection.
[063] Class IV: an extremely unstable roof W>3m. This kind of surrounding rock is extremely prone to be broken, and has an extremely large amount of rock swelling and deformation, which is often manifested as strong bulging of the two ledges, resulting in extremely poor stability of the two ledges, thereby affecting the stability of the roof. Therefore, a joint support based on the anchorage net spraying is used to support the ledge part.
[064] As a further solution of the present invention, in the step 3, the specific content of a targeted support parameter design for different types of surrounding rocks according to the classification results of the surrounding rock of the roadway is as follows:
[065] (1) Design of support parameters of the stable roof of Class I;
[066] The plastic zone of the stable roof of Class I is small, and the maximum roof fall hidden danger depth does not exceed 1.5m, which is within a range of a height of the existing anchorage rod support. Therefore, an ordinary anchorage rod is used for support. The specific support parameters are designed based on a suspension theory;
[067] a) Length design of the anchorage rod of the roof of Class I
[068] The length of the anchorage rod is calculated with reference to the following formula:
12 + 13
[069] {1 11 - 1 + 2 2 +1H
[070] Where: -an exposed length of the anchorage rod, which depends on the thickness of an anchorage rod tray and the thickness of a nut;
[071] 12 - an effective length of the anchorage rod, which is greater than the
maximum roof fall hidden danger depth to ensure that the anchorage end of the anchorage rod is outside the plastic zone;
[072] 13 - an anchorage length of the anchorage rod, which is 0.5~ 1.Om and
determined by a pull-out test;
[073] b) Design of row spacing between the anchorage rods of the stable roof of Class I
[074] The row spacing between the anchorage rods is calculated with reference to the following formula:
G = abyH
ab
[075] yH
[076] Where: G- a rock load of anchorage rod suspension. Taking into account the safety of support, the suspension load is taken as a total weight of the rock formation within the maximum roof fall hidden hazard height;
[077] ' - a bulk density of the suspended rock formation;
[078] a - the spacing of the anchorage rod
[079] b-row spacing of the anchorage rod
[080] (2) For the design of support parameters for the moderately stable roof of Class II, the unstable roof of Class III, and the extremely unstable roof of Class IV, rock formations of the moderately stable roof of Class II, the unstable roof of Class III, and the extremely unstable roof of Class IV all appear larger ranges of the plastic zone, resulting in that the roof is prone to have arch fall, and the roof fall hidden danger height exceeds 1.5m. The use of anchorage rod support alone cannot achieve the objective of controlling a malignant expansion of the plastic zone, therefore, it is necessary to use a joint support of the anchorage rod and a cable, and the specific support parameters are designed based on a falling arch theory;
[081] a) The support lengths of the anchorage rods of the roofs of Classes II, III and IV and the anchorage rod of the roof are designed with reference to the following formula:
1 =- 1 +12 + 13
12=Max(h,1.5) h =Min (R sin 0) - a
[082]
[083] where: -the exposed length of the anchorage rod, which depends on the thickness of the anchorage rod tray and the thickness of the nut; h-the minimum roof
hidden hazard height; 2 -the effective length of the anchorage rod, which is the maximum value between the minimum hidden hazard height and 1.5m, to ensure that
the anchorage end part of the anchorage rod is located outside the plastic zone; 3 -the
anchorage length of the anchorage rod is 0.5 - 1.0m, which is determined by the pull
out test;
[084] b) Design of a support length of roof anchorage cables of Classes II, III and IV
[085] The length of the anchorage cable is designed with reference to the following formula:
I =--l, + 1, + 13
[086] {2
[087] where: -the exposed length of the anchorage cable, which depends on
the thickness of the anchorage cable tray and the thickness of the nut; 2 -the effective length of the anchorage cable, which is greater than the maximum roof hidden hazard height to ensure that the end of the anchorage cable anchorage section is outside the
plastic zone to give play to the suspension function of the anchorage cable; 3-the anchorage length of the anchorage cable, which is 0.5~ 1.Om and determined by the pull-out test;
[088] c) Design of row spacing between the anchorage rods (cables) of the roofs of Classes II, III and IV
Pxc n> xk Rt PC
P=--27Ha I1---)2
[089]
[090] Where:
[091] n-the number of anchorage rods/cables in each row; P-a load of the anchorage rods/cables selected and adjusted; k-a safety factor; c-the row spacing R between anchorage rods/cables; R -the breaking load of the anchorage rod/cable; a-a half-width of the roof; b-the maximum radius of the plastic zone of the surrounding rock at a junction of the roadway roof and the ledge parts at two sides;
[092] Regarding the design of the support parameters of the ledge part, the specific support parameters are designed according to a support theory of the loose circle.
[093] As a further solution of the present invention, in step 4, the specific content of modifying the support parameters according to the on-site monitoring feedback information is as follows:
[094] Support monitoring content mainly includes the following three aspects:
[095] (1) Roof rock formation deformation, which mainly uses various roof rock formation displacement monitoring instruments to monitor the rock formation displacement in and outside a roof anchorage area;
[096] (2) A stress state of the support body such as the anchorage rod cable and so on, which uses a forcemeter of the anchorage rod cable and so on to monitor the stress state of the support body to determine the support state of the support body;
[097] (3) Observation of the surface displacement of the roadway, which uses a conventional measurement method of a cross method to monitor an approaching amount of the roof and floor of the roadway and an approaching amount of the two ledges;
[098] In on-site actual application of the forgoing three monitoring contents, one specific threshold value is set for each monitoring content. The specific threshold value is determined according to the relevant industry regulations or according to the on-site actual situation. In addition, in on-site actual monitoring, one or several contents are selected and monitored according to the needs. After monitoring data is obtained, the monitoring data are compared with specific thresholds to determine the rationality of the support, so that the monitoring data can be modified and adjusted in time.
[099] The beneficial effects of the present invention are as follows: in the present invention, the support parameters designed on the basis of the size of the plastic zone are scientific and reasonable. Based on the method for quantitatively designing the roadway support based on the size of the plastic zone of the surrounding rock, classification of the surrounding rock of the roadway is carried out by taking a radius size of the plastic zone as a main index on the basis of obtaining the radius of the plastic zone of the roadway and geometric characteristics of a boundary, so that difference support parameter design is carried out on different types of surrounding rocks so as to ensure the target and economy of roadway support design. For roadway support, the method is capable of playing roles in optimizing the support parameters of the surrounding rock and preventing the roadway from roof fall accidents.
[0100] FIG. 1 is a schematic diagram of steps of a method of the present invention;
[0101] FIG. 2 is a schematic diagram of a typical plastic zone distribution in application of specific embodiments of the present invention;
[0102] FIG. 3 is a classification diagram of a surrounding rock of a roadway in application of specific embodiment of the present invention;
[0103] FIG. 4 is a support diagram of a cable of a roadway of a surrounding rock of Class III in application of specific embodiment of the present invention;
[0104] FIG. 5 is a support diagram of a roadway of a surrounding rock of Class II in application of specific embodiments of the present invention;
[0105] The present invention will be further described below in conjunction with specific embodiments and the accompanying drawings.
[0106] As shown in FIG. 1, a method for quantitatively designing a roadway support based on a size of a plastic zone of a surrounding rock includes the following steps: step 1: carrying out calculation to determine a size of a radius R of a plastic zone of a surrounding rock of a roadway;
[0107] step 2: classifying the surrounding rock of the roadway according to a calculation result;
[0108] step 3: carrying out targeted support parameter design on different types of surrounding rocks according to a classification result of the surrounding rock of the roadway; and
[0109] step 4: modifying support parameters according to on-site monitoring feedback information.
[0110] After the excavation of the roadway, the surrounding rock produces a certain range of plastic failure under an action of concentrated stress, forming the plastic zone of the surrounding rock of the roadway. At the same time, after being affected by mining, a damage range of the surrounding rock is further expanded, and at the same time, the surrounding rock in the plastic zone is severely damaged and produces huge an expansion pressure and strong deformation. When the support is improper, the roadway falls into a roof. Therefore, the objective of roadway support design is to control the plastic zone and prevent roof fall.
[0111] At present, in rock mechanics and a roadway pressure theory, the calculation of a radius of the plastic zone of the surrounding rock of the roadway mainly adopts Castenay formula or Fenner formula, but the forgoing two formulas can only calculate the radius of the plastic zone of the surrounding rock of the roadway under bidirectional equal pressure conditions. . In actual engineering practice, most roadways are not in a bidirectional equal pressure stress field, so the calculation formula for the plastic zone of the surrounding rock under a bidirectional unequal pressure stress field is obtained to provide basic data for support design.
[0112] A method for calculating the radius R of the plastic zone of the surrounding rock of the roadway in the step 1 is as follows: for a circular roadway, relevant parameters are substituted according to the following formula (12) and formula (13) to directly calculate the radius R of the plastic zone. When the shape of a section of the roadway is rectangular, arched or the like, the existing technology FLAC3D computer numerical simulation software is used to obtain the radius R of the plastic zone.
[0113] A stress solution of the circular roadway of the surrounding rock under the bidirectional unequal pressure stress field is as follows;
af =. (1a 1 (A1 1-4!1+3 1"cos 20 ~7H~ 2 (+ A +- -- ) 1+3 cos 20 V2
r [(A-) 1+2 a-3 sin 26
[0114] (1)
[0115] where, r is a radial stress at any point; is a hoop stress at any point;
re is a shear stress at any point; y is a bulk density of the rock; H is a buried depth
of the roadway; ) is a lateral pressure coefficient; a is radii r of a circular roadway, a polar coordinate of any point;
[0116] In order to obtain a principal stress at a certain point in the surrounding rock of the circular roadway, a stress formula under a polar coordinate system needs to be transformed into a stress formula under a rectangular coordinate system using the following formula:
2 +F -cr 8 cos 20- rsin 20 2
cry +a a, -CO cos 20±+ rsin 20 2 2 rl Osn2 ,r cos 20 2
[0117] (2)
[0118] A calculation formula of the principal stress in elasticity is as follows:
2 2
2 2
[0119] ;
[0120] After substituting formula (1) into formula (2) and then substituting formula (1) into formula (3), the principal stress at any point in the surrounding rock of the circular roadway expressed in polar coordinates is obtained as follows:
,07 +r±9 ± q,)'+4 r2 2 2 "
cr1U +a 1' ~r c) ,2
[0121]
[0122] A stress state of one point has three main directions. When the stress state of one point is expressed by the principal stress, the stress can be expressed as:
[0123] (5)
[0124] where Si is a main deviation stress, and u-mis an average stress;
'0 +U0 +G7
[0125] (6)
[0126] When studying the stress of the surrounding rock of the roadway, the stress is simplified as a plane problem. Therefore, according to formula (5) and formula (6), a calculation formula of a principal deviation stress is obtained:
3 S=
[0127] (7)
[0128] Formula (4) is substituted into formula (7) to obtain:
sI+ 7 ++ 6 2 (8)
S r +u ________+____',
6 2
[0129]
_ Ur + (T
[0130] 3 (9)
[0131] According to a ultimate principal stress at any point and combined with Mohr-Coulomb strength criterion, the Coulomb-Mohr strength criterion expressed by the ultimate principal stress c1 and G3 is obtained, which is ultimate equilibrium condition: a 1 =2C Cos(O + +osif
[0132] 1 Sinr I3-SIn (10)
[0133] Where C is a cohesive force of an elastoplastic medium; p is an internal friction angle of the elastoplastic medium; after transforming the formula (10), the following is obtained:
[0134] (2n, -c,)+(2cr -a,)-(ot +a,)sin0=2C COS (p
[0135] After further transforming the above formula, the following is obtained:
[0136] 3 s3,S-(c,+or)sino=2Ccosp
[0137] r (11)
[0138] Therefore, based on formula (11), the following is obtained:
r ~ 2a'yH sin p cos 20{l1- 2)
[0139][0139] 2C cos#+yH(l+A)sinip-3('s, +S3) (2 (12)
[0140] Formula (12) is the solution of a radius of the minimum plastic zone of the surrounding rock of the circular roadway when the lateral pressure coefficient k is not equal to 1;
[0141] When the lateral pressure coefficient %=1, the formula (11) becomes:
[0142] 1 +s3 ) = 0 2Ccosp+±27YHsinp-3(S
[0143] Formula (9) is substituted into the forgoing formula, the following is obtained:
,YH r =a
[0144] 2CO + 27H sin (p (13)
[0145] Formula (13) is the solution of the radius of the plastic zone of the surrounding rock of the circular roadway when a lateral pressure coefficient is equal to 1.
[0146] In step 2, a method for classifying the surrounding rock of the roadway is as follows:
[0147] After obtaining the radius R of the plastic zone of the surrounding rock, using the following formula to calculate the maximum hidden danger depth of a rock formation of the roof of the roadway:
H = Max(R e sin 0) - a (14)
[0148] it ]
[0149] where, 0 is an angle between the radius of the largest plastic zone of the surrounding rock and a horizontal direction; and a is a vertical distance between a center of the roadway and a surface of the roof;
[0150] After obtaining the maximum roof fall hidden hazard depth H, the roof is divided into four categories according to a size of H:
[0151] Class I: a stable roof H<1.5m, this kind of rock formation of the roof has the smallest plastic zone range, and the maximum roof fall hidden hazard height does not exceed 1.5m. A main control object of the support design is the broken dangerous rock block;
[0152] Class II: a moderately stable roof 1.5m < H < 3.5m. This kind of rock
formation of the roof has the larger plastic zone range, and the maximum roof fall hidden hazard height does not exceed 3.5m. The roof produces a certain amount of deformation, and a main damage mode of the rock formation of the roof is a beam-type span, so a main control object of a support design is an unstable rock formation with the beam-type span on the roof;
[0153] Class III: an unstable roof 3.5m < H < 5.5m. This kind of rock formation of
the rock has the larger plastic zone range, and the maximum roof fall hidden hazard height does not exceed 5.5m. The roof usually undergoes a large deformation, resulting in a certain amount of roof collapse. A main object of the support design is a cracked rock that is easy to fall;
[0154] Class IV: an extremely unstable roof H>5.5m. This kind of rock formation of the roof has the largest plastic zone, and the maximum roof fall hidden hazard height exceeds 5.5m. This is mainly due to the large nonlinear deformation of the surrounding rock, which makes the roof prone to arched caving. So the main control object of the support design is a broken rock under a vault;
[0155] Similarly, the maximum radius of the plastic zone of the ledge part of the surrounding rock of the roadway is obtained according to the following formula:
W= Max (ResinO)-w Y. IT (15)
[0156]
[0157] where: w-a half width of the roadway;
[0158] After obtaining the maximum radius W of the plastic zone of the ledge part, combined with a support theory of the loose circle of the surrounding rock in the prior art, the surrounding rock of the ledge part of the roadway is divided into four categories:
[0159] Class I: a stable surrounding rock W<0.4m. The surrounding rock has a small swelling deformation. Generally, this kind of surrounding rock does not need anchorage rod support, and is supported by auxiliary support structures such as naked or steel mesh;
[0160] Class II: a moderately stable surrounding rock 0.4m < H < 1.5m. This kind
of surrounding rock has a large amount of fracture and deformation, resulting in poor stability on both sides of the roadway. Therefore, a conventional anchorage rod is used for support;
[0161] Class III: a unstable roof 1.5m < H < 3m. This kind of surrounding rock has
a wide crushing range, a large crushing depth, and a large deformation of the surrounding rock, resulting in obvious appearance of a bulge of the ledge part, poor stability of the two ledges, and anchorage net spraying is required for support and protection.
[0162] Class IV: an extremely unstable roof H>3m. this kind of surrounding rock is extremely prone to be broken, and has an extremely large amount of rock swelling and deformation, which is often manifested as strong bulging of the two ledges, resulting in extremely poor stability of the two ledges, thereby affecting the stability of the roof. Therefore, a joint support based on the anchorage net spraying is used to support the ledge part.
[0163] In the step 3, the specific content of a targeted support parameter design for different types of surrounding rocks according to the classification results of the surrounding rock of the roadway is as follows:
[0164] (1) Design of support parameters of the stable roof of Class I;
[0165] The plastic zone of the stable roof of Class I is small, and the maximum roof fall hidden danger depth does not exceed 1.5m, which is within a range of a height of the existing anchorage rod support. Therefore, an ordinary anchorage rod is used for support. The specific support parameters are designed based on a suspension theory; a) length design of the anchorage rod of the roof of Class I
[0166] The length of the anchorage rod is calculated with reference to the following formula:
1 - =1 +12 +13
[0167] L L 1 12 *
[0168] Where: -an exposed length of the anchorage rod, which depends on the thickness of an anchorage rod tray and the thickness of a nut;
[0169] 12 - an effective length of the anchorage rod, which is greater than the
maximum roof fall hidden danger depth to ensure that the anchorage end of the anchorage rod is outside the plastic zone;
[0170] 3 - an anchorage length of the anchorage rod, which is 0.5 - 1.Om and
determined by a pull-out test;
[0171] b) Design of row spacing between the anchorage rods of the stable roof of Class I
[0172] The row spacing between anchorage rods is calculated with reference to the following formula:
G=ahyH
ab<G
[0173] L yH
[0174] Where: G-rock load of the anchorage rod suspension, taking into account the safety of support, the suspension load is taken as the total weight of the rock formation within the maximum roof fall hidden hazard height;
[0175] I - a bulk density of the suspended rock formation;
[0176] a - spacing between the anchorage rods
[0177] b-row spacing between the anchorage rods
[0178] (2) For the design of support parameters for the moderately stable roof of Class II, the unstable roof of Class III, and the extremely unstable roof of Class IV, rock formations of the moderately stable roof of Class II, the unstable roof of Class III, and the extremely unstable roof of Class IV all appear larger ranges of the plastic zone, resulting in that the roof is prone to have arch fall, and the roof fall hidden danger height exceeds 1.5m. The use of anchorage rod support alone cannot achieve the objective of controlling a malignant expansion of the plastic zone, therefore, it is necessary to use a joint support of the anchorage rod and a cable, and the specific support parameters are designed based on a falling arch theory;
[0179] a) The support lengths of the anchorage rods of the roofs of Classes II, III and IV and the anchorage rod of the roof are designed with reference to the following formula:
1 = 1 +12 +1
12=Max(h,1.5) h = Min (R sin 0) - a
[0180]
[0181] where: -the exposed length of the anchorage rod, which depends on the thickness of the anchorage rod tray and the thickness of the nut; h-the minimum roof
hidden hazard height; 2 -the effective length of the anchorage rod, which is the maximum value between the minimum hidden hazard height and 1.5m, to ensure that
the anchorage end part of the anchorage rod is located outside the plastic zone; 13 -the
anchorage length of the anchorage rod is 0.5 - 1.0m, which is determined by the pull
out test;
[0182] b) Design of a support length of roof anchorage cables of Classes II, III and IV
[0183] The length of the anchorage cable is designed with reference to the following formula:
[ 1 11 + 12 +3
[0184] 12;
[0185] where: t-the exposed length of the anchorage cable, which depends on
the thickness of the anchorage cable tray and the thickness of the nut; 2-the effective length of the anchorage cable, which is greater than the maximum roof hidden hazard height to ensure that the end of the anchorage cable anchorage section is outside the
plastic zone to give play to the suspension function of the anchorage cable; 1-the
anchorage length of the anchorage cable, which is 0.5- ~.Om and determined by the
pull-out test;
[0186] c) Design of row spacing between the anchorage rods (cables) of the roofs of Classes II, III and IV
Pxc n> xk Rt PXC
P=--27Ha I -a)2
[0187]
[0188] Where:
[0189] n-the number of anchorage rods/cables in each row; P-a load of the anchorage rods/cables selected and adjusted; k-a safety factor; c-the row spacing
R between anchorage rods/cables;Rl -the breaking load of the anchorage rod/cable; a-a half-width of the roof; b-the maximum radius of the plastic zone of the surrounding rock at a junction of the roadway roof and the ledge parts at two sides;
[0190] Regarding the design of the support parameters of the ledge part, the specific support parameters are designed according to a support theory of the loose circle.
[0191] In step 4, the specific content of modifying the support parameters according to the on-site monitoring feedback information is as follows:
[0192] Support monitoring content mainly includes the following three aspects:
[0193] (1) Roof rock formation deformation, which mainly uses various roof rock formation displacement monitoring instruments to monitor the rock formation displacement in and outside a roof anchorage area;
[0194] (2) A stress state of the support body such as the anchorage rod cable and so on, which uses a forcemeter of the anchorage rod cable and so on to monitor the stress state of the support body to determine the support state of the support body;
[0195] (3) Observation of the surface displacement of the roadway, which uses a conventional measurement method of a cross method to monitor an approaching amount of the roof and floor of the roadway and a approaching amount of the two ledges;
[0196] In on-site actual application of the forgoing three monitoring contents, one specific threshold value is set for each monitoring content. The specific threshold value is determined according to the relevant industry regulations or according to the on-site actual situation. In addition, in on-site actual monitoring, one or several contents are selected and monitored according to the needs. After monitoring data is obtained, the monitoring data are compared with specific thresholds to determine the rationality of the support, so that the monitoring data can be modified and adjusted in time.
[0197] Specific embodiments and field applications are provided;
[0198] A return air hole in a lower crossheading of a 11030 working face of a certain mine is a circular roadway connecting a return air crossheading of the working face and a main return air way of a tray area. A section radius is designed as 2.4m and a buried depth is designed as about 700m. The surrounding rock of the roadway is mainly sandy mudstone with an average thickness of 6m. According to the method for designing the support in the present invention, after geomechanical parameters are measured, the radii of the plastic zones of the surrounding rocks of the different geological conditions of the return air hole are calculated to obtain the distribution map of the plastic zone of the roadway. The typical plastic zone distribution is shown in FIG. 2.
[0199] After obtaining the size of the plastic zone, the surrounding rock of the roadway is classified, as shown in FIG. 3. It can be seen from the classification diagram that the return air hole is mainly of Class III, accounting for 64% of the total length of the roadway, and the remaining 36% are of Class II. Based on this, initial support parameters of the roadway are obtained according to the table, and the support parameters of the return air hole are determined by referring to the support parameters of the 11030 working face by using an engineering analogy method. The row spacing between the anchorage rods of the roof of Class III is 800x900mm, a row distance between anchorage cables is 1600x1800mm, a row distance between the anchorage rods in the ledge part is 800x800mm; a row distance between the anchorage rods of the roof of Class II is 800x900mm, and the row distance between the anchorage cables is 2000x2200mm. as shown in FIGS 4 and 5.
[0200] In the present invention, based on the method for quantitatively designing the roadway support based on the size of the plastic zone of the surrounding rock, classification of the surrounding rock of the roadway is carried out by taking a radius size of the plastic zone as a main index on the basis of obtaining the radius of the plastic zone of the roadway and geometric characteristics of a boundary, so that difference support parameter design is carried out on different types of surrounding rocks so as to ensure the target and economy of roadway support design. For roadway support, the method is capable of playing roles in controlling the surrounding rock and maintain the stability of the surrounding rock, and preventing the roadway from roof fall accidents; the roadway support effect is good and the safety is high.
[0201] Although the invention has been described with reference to specific examples, it will be appreciated by those skilled in the art that the invention may be embodied in many other forms, in keeping with the broad principles and the spirit of the invention described herein.
[0202] The present invention and the described embodiments specifically include the best method known to the applicant of performing the invention. The present invention and the described preferred embodiments specifically include at least one feature that is industrially applicable
Claims (5)
1. A method for quantitatively designing a roadway support based on a size of a plastic zone of a surrounding rock, comprising the following steps:
step 1: carrying out calculation to determine a size of a radius R of a plastic zone of a surrounding rock of a roadway;
step 2: classifying the surrounding rock of the roadway according to a calculation result;
step 3: carrying out targeted support parameter design on different types of surrounding rocks according to a classification result of the surrounding rock of the roadway; and
step 4: modifying and adjusting support parameters according to on-site monitoring feedback information.
2. The method for quantitatively designing the roadway support based on the size of the plastic zone of the surrounding rock according to claim 1, wherein a method for calculating the radius R of the plastic zone of the surrounding rock of the roadway in the step 1 is as follows: for a circular roadway, relevant parameters are substituted according to the following formula (12) and formula (13) to directly calculate the radius R of the plastic zone, when the shape of a section of the roadway is rectangular, arched or the like, FLAC3D, UDEC and other computer numerical simulation methods are used to obtain the radius R of the plastic zone ;
a stress solution of the circular roadway of the surrounding rock under the bidirectional unequal pressure stress field is obtained based on the following formula;
S= (]+^) I- a+(A-) 1-4a+3a cos 20 rr2 r
Ce = (A1 -( A+ -1 r2' +3 -,cos 20 2 rr
r=L- (;t-1) 1(2 -3 sin 20 2 r r4
where, U, is a radial stress at any point; CFis a hoop stress at any point; 7T,
is a shear stress at any point; y is a bulk density of the rock; H is a buried depth of the
roadway; ? is a lateral pressure coefficient; a is radius r of a circular roadway, 9a polar coordinate of any point;
in order to obtain a principal stress at a certain point in the surrounding rock of the circular roadway, a stress formula under a polar coordinate system needs to be transformed into a stress formula under a rectangular coordinate system using the following formula:
U U+F + Icos 2 -r,, sin 2Q 2 2
= _ 2 cos 20 +rsin20 2 2 r =) ' sin20+r cos20 2 (2)
a calculation formula of the principal stress in elasticity is as follows:
Ox +Oy+ (r+4 )
IT =yr
2 2 (3) after substituting formula (1) into formula (2) and then substituting formula (1) into formula (3), the principal stress at any point in the surrounding rock of the circular roadway expressed in polar coordinates is obtained as follows: ar +a CF 1 2 2
2 2 (4)
a stress state of one point has three main directions, when the stress state of one point is expressed by the principal stress, the stress can be expressed as:
(O" =0" + S,. (5)
where Si is a main deviation stress, and am is an average stress;
=c7] + (T2 +U7 3 3 (6)
when studying the stress of the surrounding rock of the roadway, the stress is simplified as a plane problem, therefore, according to formula (5) and formula (6), a calculation formula of a principal deviation stress is obtained:
3 =3 2c~3 - u
(7)
Formula (4) is substituted into formula (7) to obtain:
- ,+ta3 (o,-ao )2+4r 2 6 2
6 2
S+S 3 _ r +
3 (9)
According to a ultimate principal stress at any point and combined with Mohr Coulomb strength criterion, the Coulomb-Mohr strength criterion expressed by the ultimate principal stress a] and a3 is obtained, which is ultimate equilibrium condition:
2 CCos p I+ sin rp 1- Sin (p I -Sin rp (10)
where C is a cohesive force of an elastoplastic medium; <p is an internal friction angle of the elastoplastic medium; after transforming the formula (10), the following is obtained:
(2a, - )i+(2 ,- o)-(aj + 3 )sin =W2Ccoso
after further transforming the above formula, the following is obtained:
3s +3s, -(C 1 ±c+ )sin P =2C cos p
a 2 7 H(1-A)sinpcos2O- =2Ccosp+yH(1+-)sin9-3(s, )
r (11)
therefore, based on formula (11), the following is obtained:
F 2a yH sin pcos29(1-2) 2Ccosf+yH(l+A)sin 9 -3(s, +s3 (12)
) formula (12) is the solution of a radius of the minimum plastic zone of the surrounding rock of the circular roadway when the lateral pressure coefficient ) is not equal to 1;
when the lateral pressure coefficient k=1, the formula (11) becomes:
2Ccos +2YH sin(p-3(s 1 +s 3 )=0
formula (9) is substituted into the forgoing formula, the following is obtained:
yH ra (13)
formula (13) is the solution of the radius of the plastic zone of the surrounding rock of the circular roadway when the lateral pressure coefficient is equal to 1.
3. The method for quantitatively designing the roadway support based on the size of the plastic zone of the surrounding rock according to claim 1, wherein in step 2, a method of classifying the surrounding rock of the roadway is as follows:
after obtaining the radius R of the plastic zone of the surrounding rock, using the following formula to calculate the maximum hidden danger depth of a rock formation of the roof of the roadway:
H=Max(ResinO)-a (14) Oe[Or]
where, 0 is an angle between the radius of the largest plastic zone of the surrounding rock and a horizontal direction; and a is a vertical distance between a center of the roadway and a surface of the roof; after obtaining the maximum roof fall hidden hazard depth H, the roof is divided into four categories according to a size of H:
Class I: a stable roof H<1.5m, this kind of rock formation of the roof has the smallest plastic zone range, and the maximum roof fall hidden hazard height does not exceed 1.5m, a main control object of the support design is the broken dangerous rock block;
Class II: a moderately stable roof 1.5m < H < 3.5m, this kind of rock formation
of the roof has the larger plastic zone range, and the maximum roof fall hidden hazard height does not exceed 3.5m, the roof produces a certain amount of deformation, and a main damage mode of the rock formation of the roof is a beam-type span, so a main control object of a support design is an unstable rock formation with the beam-type span on the roof;
Class III: an unstable roof 3.5m < H < 5.5m, this kind of rock formation of the
rock has the larger plastic zone range, and the maximum roof fall hidden hazard height does not exceed 5.5m, the roof usually undergoes a large deformation, resulting in a certain amount of roof collapse, a main object of the support design is a cracked rock that is easy to fall;
Class IV: an extremely unstable roof H>5.5m, this kind of rock formation of the roof has the largest plastic zone, and the maximum roof fall hidden hazard height exceeds 5.5m, this is mainly due to the large nonlinear deformation of the surrounding rock, which makes the roof prone to arched caving, so the main control object of the support design is a broken rock under a vault;
similarly, the maximum radius of the plastic zone of the ledge part of the surrounding rock of the roadway is obtained according to the following formula:
W = Max (R e sin )-w (1
where: w-a half width of the roadway; after obtaining the maximum radius W of the plastic zone of the ledge part, combined with a support theory of the loose circle of the surrounding rock in the prior art, the surrounding rock of the ledge part of the roadway is divided into four categories:
Class I: a stable surrounding rock W<0.4m, the surrounding rock has a small swelling deformation, generally, this kind of surrounding rock does not need anchorage rod support, and is supported by materials of auxiliary support structures such as naked or steel mesh;
Class II: a moderately stable surrounding rock 0.4m < H < 1.5m, this kind of
surrounding rock has a large amount of fracture and deformation, resulting in poor stability on both sides of the roadway, therefore, a conventional anchorage rod is used for support;
Class III: a unstable roof 1.5m < H < 3m, this kind of surrounding rock has a
wide crushing range, a large crushing depth, and a large deformation of the surrounding rock, resulting in obvious appearance of a bulge of the ledge part, poor stability of the two ledges, and anchorage net spraying is required for support and protection.
Class IV: an extremely unstable roof H>3m, this kind of surrounding rock is extremely prone to be broken, and has an extremely large amount of rock swelling and deformation, which is often manifested as strong bulging of the two ledges, resulting in extremely poor stability of the two ledges, thereby affecting the stability of the roof, therefore, a joint support based on the anchorage net spraying is used to support the ledge part.
4. The method for quantitatively designing the roadway support based on the size of the plastic zone of the surrounding rock according to claim 1, wherein in step 3, the specific content of target parameter design for different types of surrounding rock according to the classification results of the surrounding rock of the roadway is as follows:
(1) design of support parameters of the stable roof of Class I;
the plastic zone of the stable roof of Class I is small, and the maximum roof fall hidden danger depth does not exceed 1.5m, which is within a range of a height of the existing anchorage rod support, therefore, an ordinary anchorage rod is used for support, the specific support parameters are designed based on a suspension theory; a) length design of the anchorage rod of the roof of Class I the length of the anchorage rod is calculated with reference to the following formula:
I= II + , + 13 12 3
Where: -an exposed length of the anchorage rod, which depends on the thickness of an anchorage rod tray and the thickness of a nut;
2 - an effective length of the anchorage rod, which is greater than the
maximum roof fall hidden danger depth to ensure that the anchorage end of the anchorage rod is outside the plastic zone;
3 - an anchorage length of the anchorage rod, which is 0.5 - 1.Om and
determined by a pull-out test;
b) design of row spacing between the anchorage rods of the stable roof of Class
the row spacing between the anchorage rods is calculated with reference to the following formula:
yH where: G- a rock load of anchorage rod suspension, taking into account the safety of support, the suspension load is taken as a total weight of the rock formation within the maximum roof fall hidden hazard height;
- a bulk density of the suspended rock formation;
a - spacing between the anchorage rods
b-row spacing between the anchorage rods
(2) For the design of support parameters for the moderately stable roof of Class II, the unstable roof of Class III, and the extremely unstable roof of Class IV, rock formations of the moderately stable roof of Class II, the unstable roof of Class III, and the extremely unstable roof of Class IV all appear larger ranges of the plastic zone, resulting in that the roof is prone to have arch fall, and the roof fall hidden danger height exceeds 1.5m. The use of anchorage rod support alone cannot achieve the objective of controlling a malignant expansion of the plastic zone, therefore, it is necessary to use a joint support of the anchorage rod and a cable, and the specific support parameters are designed based on a falling arch theory;
a) the support lengths of the anchorage rods of the roofs of Classes II, III and IV and the anchorage rod of the roof are designed with reference to the following formula:
I =1,+1, + 1
12=Max(h,1.5) h = Min (RsinO)-a
where: -the exposed length of the anchorage rod, which depends on the thickness of the anchorage rod tray and the thickness of the nut; h-the minimum roof
hidden hazard height; 2 -the effective length of the anchorage rod, which is the maximum value between the minimum hidden hazard height and 1.5m, to ensure that the anchorage end part of the anchorage rod is located outside the plastic zone; 3 -the anchorage length of the anchorage rod is 0.5 ~ 1.0m, which is determined by the pull out test; b) design of a support length of roof anchorage cables of Classes II, III and IV the length of the anchorage cable is designed with reference to the following formula:
I= 1 +2 +13 12 >H
where: -the exposed length of the anchorage cable, which depends on the
thickness of the anchorage cable tray and the thickness of the nut; 2-the effective length of the anchorage cable, which is greater than the maximum roof hidden hazard height to ensure that the end of the anchorage cable anchorage section is outside the
plastic zone to give play to the suspension function of the anchorage cable; 3 -the anchorage length of the anchorage cable, which is 0.5 1.Om and determined by the
pull-out test;
c) design of row spacing between the anchorage rods (cables) of the roofs of Classes II, III and IV
Pxc n ;>xk R PXC I a)2 P=27Ha[1- (-)
where: n-the number of anchorage rods/cables in each row; P-a load of the anchorage rods/cables selected and adjusted; k-a safety factor; c-the row spacing between anchorage rods/cables;Rl -the breaking load of the anchorage rod/cable; a-a half-width of the roof; b-the maximum radius of the plastic zone of the surrounding rock at a junction of the roadway roof and the ledge parts at two sides; regarding the design of the support parameters of the ledge part, the specific support parameters are designed according to a support theory of the loose circle.
5. The method for quantitatively designing the roadway support based on the size of the plastic zone of the surrounding rock according to claim 1, wherein in step 4, the specific content of the support parameters modified according to the on-site monitoring feedback information is as follows:
support monitoring content mainly includes the following three aspects:
(1) roof rock formation deformation, which mainly uses various roof rock formation displacement monitoring instruments to monitor the rock formation displacement in and outside a roof anchorage area;
(2) a stress state of the support body such as the anchorage rod cable and so on, which uses a forcemeter of the anchorage rod cable and so on to monitor the stress state of the support body to determine the support state of the support body;
(3) observation of the surface displacement of the roadway, which uses a conventional measurement method of a cross method to monitor an approaching amount of the roof and floor of the roadway and a approaching amount of the two ledges;
in on-site actual application of the forgoing three monitoring contents, one specific threshold value is set for each monitoring content, the specific threshold value is determined according to the relevant industry regulations or according to the on-site actual situation, in addition, in on-site actual monitoring, one or several contents are selected and monitored according to the needs, after monitoring data is obtained, the monitoring data are compared with specific thresholds to determine the rationality of the support, so that the monitoring data can be modified and adjusted in time.
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| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN113062773A (en) * | 2021-04-16 | 2021-07-02 | 中国人民解放军军事科学院国防工程研究院工程防护研究所 | Automatic monitoring and early warning system and construction method for deep tunnel assembly type ventilation vertical shaft |
| CN113863950A (en) * | 2021-09-18 | 2021-12-31 | 中煤科工开采研究院有限公司 | Roadway arrangement method for reserving small coal pillars between stope faces |
| CN114233390A (en) * | 2021-12-20 | 2022-03-25 | 中铁二院工程集团有限责任公司 | Auxiliary tunnel plugging structure suitable for weak surrounding rock and calculation method thereof |
| CN114547810A (en) * | 2022-04-25 | 2022-05-27 | 中国矿业大学(北京) | High-prestress energy absorption control design method for coal mine dynamic disaster |
-
2021
- 2021-01-22 AU AU2021100425A patent/AU2021100425A4/en not_active Ceased
Cited By (6)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN113062773A (en) * | 2021-04-16 | 2021-07-02 | 中国人民解放军军事科学院国防工程研究院工程防护研究所 | Automatic monitoring and early warning system and construction method for deep tunnel assembly type ventilation vertical shaft |
| CN113863950A (en) * | 2021-09-18 | 2021-12-31 | 中煤科工开采研究院有限公司 | Roadway arrangement method for reserving small coal pillars between stope faces |
| CN114233390A (en) * | 2021-12-20 | 2022-03-25 | 中铁二院工程集团有限责任公司 | Auxiliary tunnel plugging structure suitable for weak surrounding rock and calculation method thereof |
| CN114233390B (en) * | 2021-12-20 | 2023-07-25 | 中铁二院工程集团有限责任公司 | Auxiliary gallery plugging structure suitable for weak surrounding rock and calculation method thereof |
| CN114547810A (en) * | 2022-04-25 | 2022-05-27 | 中国矿业大学(北京) | High-prestress energy absorption control design method for coal mine dynamic disaster |
| CN114547810B (en) * | 2022-04-25 | 2022-07-15 | 中国矿业大学(北京) | High-prestress energy absorption control design method for coal mine dynamic disaster |
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