Abstract Liling Mine of Huaibei Coal Mining Administration is a gassy and o utburst coal mine. The coalbed methane reserves in Luling Mine is up t o 6.4 billion mª¬3 and the permeability of the seams are very low. At present, the gas drainage in the mine is dominated by cross-measure bo reholes in roof and the max. gas drainage efficiency from a working fa ce is 20 to 25 %. Therefore, this paper has analyzed the gas release a nd relaxation zones in the coal mining face of Luling Mine and provide d with four improved drainage methods, such as cross-measure boreholes in floor, long horizontal boreholes drilled in roof strata, cross-me asure boreholes in the roof and surface vertical bore hole. A comparis on of the four options was conducted. The results show that an integra ted approach with the cross-measure boreholes in the floor and surfac e vertical boreholes can drain gas from a large area with high efficie ncy. 1 Introduction Luling Mine is a high gas mine with an outburst potential. The min e is located 20 km southeast of Shuozhou City. The original designed c apacity of the mine was 1ª±5 million t/y. In 1993, the mine was expand ed to a design capacity of 2ª±4 million t/y. In 1994, its annual prod uction was 1ª±85 million t and the mine had become a major mine of the Huaibei Coa Mining Administration. There is a recoverable coal reserves of 100 million tons to a dept h of 700m at the mine and there is an additional reliably estimated c oal reserves of 271 million tons between the depth of 700 m and 1200 m at the mine. The whole mine field is covered by Quaternary strata with a thickn ess of 129 to 237 m. The coal - bearing formation is mainly Permian Sh ihezi and Shanxi Group with a total of 8 seams. Among them, there are 4 main minable seams and they are No. 7, No. 8, No. 9 and No. 10 seams with a dip angle of 8ª¬o to 22ª¬o. The coal belongs to 1/3 coking co al and gas fat coal. The mine uses a staged cross cut development and mine shaft. And t he mine is divided into three levels, iª±e., -400m level, -590m level and -700m level. At present, the mining panel in the middle part of th e first level has been mined out and the mining operations are in the second level. Mining method of the mine is an inclined slicing and lon gwall mining. A conventional mechanized coal mining method is mainly u sed and a full caving method is used for the roof control. No. 8 and No. 9 seams are highly gassy seams with an outburst pote ntial. There were 20 outbursts occurred up to the present. One of the largest outbursts was 314 t. The No. 8 and No. 9 seams are also very t hick seams with a low permeability. According to the measurements, the permeability coefficients of the two seams are 0ª±067 md and the two seams are very difficult in gas drainage. The gas drainage radius of the two seams are only 0ª±8 m. The gas content and gas pressure in dif ferent levels of each seam are shown in Table 1. The gas pressure grad ient was 0ª±011 MPa/m. The gas reserves above -1200 m is 6ª±4 billion mª¬3. Table 1 Table of gas pressure and gas content in different level Level Gas pressure (kg/cmª¬2) Gas content in each seam (mª¬3/t) Noª±7 seam Noª±8 seam Noª±9 seam -400 23.5 14.35 16.22 15.25 -440 26.88 15.06 17.34 16.30 -480 30.24 15.69 18.33 17.23 -520 33.6 16.25 19.21 18.06 -560 36.7 16.9 20.2 19.0 -600 40.6 17.8 20.9 19.8 At present, the absolute gas emission of the mine is 55 mª¬3/min. and the relative emission is about 15 mª¬3/t. The absolute gas emissio n of No. 8 seam is generally about 6 to 8 mª¬3/min. and 10 to 15 mª¬3/ min. at the deeper part of the seam. Two gas drainage systems were set up at the central ventilation shaft and west ventilation shaft. The a nnual pure gas drained is up to 3ª±6 million mª¬3 and the average gas drainage is 7ª±0 mª¬3/min. The gas drained can provide 4000 households with gas fuel. In the past years, a number of underground gas drainage methods w ere used, such as cross-measure boreholes in floor, in-seam bore hole, roof horizontal bore hole in air gate, rock roadway of roof and othe rs. At present, only the gas drainage with cross-measure boreholes (th ere are 1355 holes each with average gas flow of 0ª±004 mª¬3/min. ) an d the pregrouting before the development of cross cut are used to prev ent the outburst with a good result. Other methods were not being used due to restrictions of mining and driving operation. Therefore the ga s drainage rate of the mine is normally around 15% and can not be incr eased. Additionally£¬when the first slicing face of the No. 8 seam is in operation, 60 % of the gas released from upper and lower adjacent stra ta will come to the first slicing face and with a ventilation of 800 m ª¬3/min., the gas content in the face is often over the limit. So prod uction at the face will have to be halted from time to time. From the above, it can be seen that with increasing mining depth, the gas control issue at Luling Mine is far from solved. Meanwhile, th e mine annually discharges 30 million mª¬3 of methane to the atmospher e. The discharge will not only pollute the air, but waste valuable re sources. Therefore, based on the previous study and according to actual pra ctices and the improved technology, 4 technical options are proposed. The author believes that once the options can be applied, it will be p ossible to drain the released gas from No.and No. 9 seams from a large area which can basically solve the difficult problem of gas vestricti ng safe Production. 2 Post-mining Relaxation and ¡°O¡± Ring Channel It is well known that fractures caused by overburden displacement (as shown in Fig. 1) after mining can be divided into ¡°three vertical zones¡± and ¡°three latteral areas¡±. That¡äs to say, along the verti cal direction, from bottom to top, the goaf can be divided into a cav ing zone, fracture zone and bending subsidence zone, along the advanci ng direction of the coal face, the goaf can divided into coal wall su pported area, bed separation area and recompressed area. Meanwhile, th e mining induction also will damage the coal face floor, causing ¡°thr ee bottom zones¡± and ¡°three bottom areas¡±. From the top to the bott om of the floor in the goaf, the floor can be divided into a direct fa ilure zone, affected zone and slight deformation zone. Along the advan cing direction of the coal face, the floor can be divided into a suppo rt area, bed separation area and compressed area. With the advancing o f the face, ¡°the three zones¡± and ¡°three areas¡± in the roof and fl oor of the face will be advanced also. This space and dynamic concepti on is very important to the seam pressure releasing and gas flow.Fig. 1 ¡°Vertical three zones¡± and ¡°Horizontal three areas¡± A - coal wall support area ( a-b ), B - bed separation area ( b-c ), C - recompressed area ( c-d ), I - caving zone; II - fracture zone; II I - bending subsidence zone; a - support influence angle. In 1989, the author first introduced the concept that a mining induc ed ¡°O¡± type fracture zone ( with a width of 20 to 30 m ) in the over burden is a favorable channel for gas releasing in the goaf. The autho r also applied the channel to the gas drainage plan with cross-measure borehole in roof and surface vertical borehole. When the face roof caved, the central part of the caved material w ill be quickly compacted in the goaf and at each side of the goaf, due to the support function of the coal pillars, the fractures from bed s eparation will not be easily compacted and closed and can be permeable to each other in all directions. Thus those bed separation fractures will form a permeable ring type district which is quite similar to the ¡° O - X ¡± form of the broken main roof, called ¡°O¡± shape ring ch annel. In the recent time, asked by Luling Mine of Huaibei Coal Mining Administration, China University of Mining and Technology had conduct ed a study on simulation test of mining induced fractures in overburde n over No. 8 seam of Luling Mine and ¡°O¡± ring features and the study has further proved ¡°O¡± ring channel (Fig. 2). According to the simulation test, the first slicing face of No. 8 seam will have a mining height of 2ª±5 m and the height of mining indu ced caving zone in the overburden will be 8 to 9 m which is 3ª±2 to 3 ª±6 times the mining height. The fracture zone will have a height of 24 to 27 m which is 9ª±6 to 10ª±8times the mining height. The broken angle of the fracture will be=65ª¬o in strike,=70ª¬o in upper, r=75ª¬o in lower, and a=15ª¬o; a. A diagram of inclined cross section.b. Plane view Fig.2 A diagram of ¡°O¡± ring simulation test. the width of¡°O¡±ring channel will be 33-34 m.The tests also showed mining induced damages to the floor , and the direct failure zone will have a depth of 10 to 15 m, influenced zone is 20 to 30 m deep, and the slight deformation zone has a depth of 50 to 70 m. Meanwhile, at the sides of coal pillars in the goaf, there will be a ¡°O¡± ring chan nel in the floor. As the distance between No. 7 and No.8 seam is 20m, and about 3ª±3 m between No. 8 seam and No. 9 seam. Those seams are all in the upper crack zone and lower direct failure zone of the first slicing face of No. 8 seam. Therefore the released gas from No. 7 seam, lower slicing face of No. 8 seam and No. 9 seam all rushed into the goaf of the fir st slicing face of No. 8 seam. It is definite that the gas content in the first slicing face was over the limit. 3 Technical Options of Gas Drainage Base on the analysis of gas source and ¡°O¡± ring channel, a ratio nal placement of boreholes would result in high efficient gas drainage from a large area over a relatively long period from the goaf. Theref ore after careful consideration, four technological options were provi ded. It takes 2812-1 face as a trial face. The face conditions are: 5 00 m from a strike length of 1200 m, a face width of 120m, a return ai r way with an elevation of -390 m, the air intake way with an elevatio n of -425 m and the seam inclination of 15ª¬o. The coal reserves and g as reserves of each seam are shown in Table 2. Table 2 Gas Reserves of Each Seam Item No. 8 seam Full seam First slicing No. 9. seam No. 7 seam Total Seam thickness(m) 9 2.25 3.2 0.7 Minable reserves (1000 t) 756 189 252 58.8 Gas content(mª¬3/t) 17 16 15 Total gas reserves(1000t/mª¬3) 12852 3213 4032 882 17766 (1) Cross-measure boreholes in the roof As shown in Fig.3, before mining operation in No. 8 seam, at a cen tralized roadway and a railway under the floor of No. 9 seam, cross-m easure boreholes, totalling 5 groups were drilled to the top and low p arts of the air intake way of No. 8 seam. Each group has three borehol es, at left, middle and right, all of them are open holes. If the mi ddle hole is changed to casing holes, and the left and right holes wil l be still open holes, there will be 6 to 7 additional casing holes an d keep 8 to 9 casing holes in the downward section of air intake way. Thus there are 19 holes located in the inclined cross section, and 10 holes are open hole and 9 casing ones. All casing holes will have a 1 08 mm casing cementing for opening hole and then have a 89 mm casing down to the bottom. Screen tube is always used at section of coal sea m. Those holes were used for gas predrainage before the coal mining fa ce in operation. After the first slice of No.8 seam has been mined the open holes ceased to function the casing holes were kept for gas drai nage released from the face floor. Based on the bore hole pattern, the control area of the gas drainage could be 116 % over the mining face. In this way, its function is to prevent the gas outburst, and also co uld reduce the released gas rushing into the first slicing face of No. 8 seam. (2) Long horizontal roof borehole Horizontal roof boreholes has been used previously in Luling mine. Because the boreholes could not be fully located in ¡°O¡± ring channe l, the length of boreholes is less than 100 m. It also needed a lot o f drilling sites and engineering work. The drainage results were not s o good. If it uses a connection cross-cut to develop a raise to the roof of No. 8 seam with a distance of 10 m for a drilling site, at the raise 2 to 3 horizontal boreholes are drilled to the direction of the open-off cut with a length of 300 m. The 2 to 3 holes would be locate d in the ¡°O¡± ring channel. The bore hole structures are opening hole s with a diameter of 146 mm, a 127 mm casing pipe for cementing, and a 108 mm open hole down to the bottom of the hole. The length of a hor izontal roof borehole is dependent on a drilling equipment. At presen t, China has imported two sets of drilling rigs capable of drilling ho rizontal boreholes of 1000m long from USA. Luling Mine has purchased o ne set drilling rig developed by Xi¡äan Branch of Central Coal Mining Research Institute and the drilling rig has a horizontal drilling capa city of 300 m. In the face with a strike of 500 m, only two drilling s ites are required to drill the horizontal boreholes. Fig.3 Diagram of cross-measure floor borehole.Fig. 4 Cross sectio n of surface drilling. (3) Cross-measure roof borehole As shown in Fig.3, No. 10 borehole is cross-measure roof borehole. A drilling site was developed in connection cross-cut of an air inta ke way. The opening hole of No. 10 hole was drilled at the drilling si te with an angle of 25ª¬o to ¡°O¡± ring channel at the upper of the co al face via upper section of the goaf. The total length of the bore ho le was 75 m with a 127 mm opening hole, a 108 mm casing pipe, a 89 mm open hole to the bottom of the hole and distance between boreholes of 70 to 80 m. The gas drainage by cross-measure borehole is basicall y same as the horizontal bore holeª±The disadvantages of the ª©cross- ªª Table 3 Comparison Table for four gas drainage alternatives 1.Cross-measure floor borehole 2. Horizontal roof borehole 3.Cross-mea sure roof borehole 4.surface vertical bore hole Advantages: 1. Predrainage before mining, released gas drainage after mining, one hole two functions; 2.The gas drainage is the efficient measure to prevent gas outburst fo r mechanized roadway heading, 3.Good result in releasing gas drainage from floor. 1.Less bore holes required 2.Bore holes horizontally into ¡°O¡± ring channel, a high concentratio n gas can be continuously drained out. 1. Less bore hole required, 2. Penetrated into ¡°O¡± ring channel, high density gas can be drained . 1. Less bore hole required, 2. Penetrated through ¡°O¡± ring channel, good result of gas drainage from No. 7 seam and goaf, 3. Gas drainage with a long period. Disadvantages: 1. More bore holes required, little quantity in predrainage, 2. the drilling site is affected by mining induction, a lot of mainten ance and support for the drilling site, 3. It is difficult to keep bore hole with a long period. 1. Horizontal bore hole with a length, must have special equipment, high operation technology, 2. Raise and drilling site affected by mining induction, difficult to keep maintenance and support. 3. Gas drainage have a short period. 1. Bore hole through goaf, drilli ng operation difficult, 2. Drilling site is affected by mining induction, difficult to keep ma intenance and support. 1. Large engineering work, high cost, 2. Set up a surface gas drainage system, 3. Surface construction restricted by countryside. measure borehole were difficult for maintenance and support of the con nection cross-cut of air way and drilling site, and also it is difficu lt to operate the bore hole drilling through a goaf and it should have a grouting operation, while drilling. (4) Surface vertical borehole Before mining operation, a surface vertical borehole is drilled to the top of the mining induced ¡°O¡± ring channel, and gas can be reco vered by surface borehole, see Fig.4 and Fig.5. Because of the bore ho les being through No. 7 seam and the fracture zone of No. 8 seam, a go od result of gas drainage from goaf can be obtained. A successful gas drainage with surface vertical boreholes in 1018 coal face of Taoyuan Mine has proved that the key point is the inclination of a bore hole not over 5/1000. Fig.5 Diagram showing the placement of surface borehole (5 ) Comprehensive comparison of gas drainage options can be seen i n Table 3. 4 Prediction of Gas Drainage Options It is supposed that the followings are the prediction results of t he four gas drainage options used at 2812-1 coal face: 4.1 Prediction of gas emission from the first slicing face of No. 8 s eam. (1) Gas predrained and emitted before mining operation from the f irst slicing face of No. 8 seam. Based on the gas predrainage over two years with cross-measure bor eholes, the gas drainage rates for No. 8 seam and No. 9 seam are 25 % and 20 % respectively as predicted. The gas drainage from No. 8 seam will be 3ª±21 million mª¬3 and the gas drainage from No. 9 seam will be 0ª±8 million mª¬3. The total will be 4ª±01 million mª¬3. Based on the gas emission of 3ª±0 mª¬3/min. of the gate way headin g in the first slicing face of No. 8 seam, the gas emission will be 0 ª±955 million mª¬3. The total gas reserves of No. 8 and No. 9 seams is 16ª±884 million mª¬3. Subtracting the gas predrainage and emission quantities of 4ª±9 65 million mª¬3, the remaining gas reserves will be 11ª±919 million m ª¬3 and the gas content is 11ª±82 mª¬3/t. (2) Gas emission of the first slicing face of No.8 seam. Relative gas emission from a coal face is calculated with the foll owing formula, Qr =Qª­1+Qª­2 where: Qª­1¡ª¡ªrelative gas emission from mining seam, mª¬3/t£¬ Qª­2¡ª¡ªrelative gas emission from adjacent seam, mª¬3/t. Relative gas emission quantity of the mining seam is calculated wi th the following formula, Q=X¡ªXª­k where: X¡ª¡ªgas content of a coal seam, here is the residual gas content after predrainage and emission, Xª­k¡ª¡ªresidual gas content of a seam, here, Xª­k=3 mª¬3/t; Qª­1=11ª±82-3=8ª±82 (mª¬3/t); The relative gas emission from adjacent seam is calculated with th e following formula, Qª­2 = ¡Æmª­i mª­0(Xª­i-Xª­ª©kiªª)1-hª­i hª­p of which: mª­i¡ª¡ªThickness of adjacent seam, m; mª­o¡ª¡ªThickness of mining seam, m; Xª­i¡ª¡ªgas content of adjacent seam, mª¬3/t; Xª­ª©kiªª¡ª¡ªresidual gas content in adjacent seam, mª¬3/t; hª­i¡ª¡ªdistance between the working seam and adjacent seam, m, hª­p¡ª¡ªdistance between the working seam and the affected seam which does not emit gas to the mining seam, m. No. 9 seam is the lower adjacent seam of No. 8 seam and has a dist ance of 3ª±3 m to No. 8 seam. No. 9 seam has a thickness of 3ª±2 m and a residual gas content of 11ª±82 mª¬3/t. Because of a gently inclined seam, hp should be put in the above formula and the relative gas emis sion of No. 9 seam is 4 mª¬3/t. No.7 seam is the upper adjacent seam of No. 8 seam and has a dista nce of 20 m to No. 8 seam. The thickness of the No. 7 seam is 0ª±7 m a nd the gas content is 15 mª¬3/t. After calculation, the relative gas emission of No. 7 seam is 0.9 mª¬3/t. Qª­2=4+0ª±9=4ª±9 mª¬3/t. when No. 8 seam is mined, the total gas emission is, Qª­r=8ª±82+4ª±9=13ª±72 (mª¬3/t) No. 8 seam has a thickness of 9 m and is divided into four slicing for coal mining. Each slicing has a thickness of 2.25 m. According to the mine statistics, the percentage of the relative gas emission quan tity for No. 1, No. 2, No.3 and No.4 slicing are individually as aª­1= 60 %, aª­2=20 %, aª­3=13 % and a4=7%. The relative gas emission quanti ty of the first slicing of No. 8 seam is: Qª­r=13ª±72¡Á0ª±6¡Á=32ª±92 (mª¬3/t). Based on the average advancing rate of 1ª±1 m /d of the first slic ing face, a daily production of 415 t/d and a monthly production of 1 2450 t, the absolute gas emission quantity is , Qª­a = 415 x 32ª±92 ¡Â 1440 = 9ª±48 (mª¬3/min). The ventilation at the face must be over 1000 mª¬3/min to ensure t hat gas content is under the limit in ventilation air. 4.2 Prediction of gas drainage. According to actual information, the gas drainages are predicted a s shown in Table 4. 4.3 Evaluation of feasibility for gas drainage options.Results are sho wn in Table 5. 5 Conclusion (1)Due to a low permeability of No. 8 and No. 9 seams, with cross- measure floor boreholes as a main method for gas drainage, the maximum . gas drainage rate of the coal face was 20 to 25 % which could not so lve the problems of gas content over rated level and discontinued prod uction in the first slicing face of No. 8 seam. (2)The mining induced ¡°O¡± ring channel in the roof and floor is the favorable condition for goaf gas drainage. Base on this condition, three options with different combinations of four kinds of boreholes can result in gas drianage over a large area and over a long period of time have a large area which could not only solve the outburst in coa l roadway, but also could prevent gas content over the rated level in the first slicing face of No. 8 seam in order to create conditions for mine production and development. Table 4 Prediction of gas drainage for the four gas drainage metho ds. Method of gas drainage Drainage quantity from one hole (mª¬3/min.) Num ber of normal gas drainage ( hole ) Total gas quantity drained (mª¬3/m in.) (1)Cross-measure floor casing borehole. (2)Horizontal roof borehole (3)Cross-measure roof borehole. 4.surface vertical bore hole. 0.25 1.0 1.0 1.0 9 2 2 2 2.25 2.0 2.0 3.0 Table 5 Feasibility evaluation table for combined gas drainage options Combined gas drainage method Predicted gas drainage quantity (mª¬3/min.) Gas emission quantity before gas drainage (mª¬3/min.) Gas emission quantity after gas drainage (mª¬3/min.) Ventilation quantityat coal face (mª¬3/min.) Gas content in air returning way (mª¬3/min.) Evaluation (1) + (2) (1) + (3) (1) + (4) 2.25 + 2= 4.25 2.25 + 2= 4.25 2.25 + 3= 5.25 9.48 9.48 9.48 5.23 5.23 4.23 600 600 600 0.87 0.87 0.70 Feasible Feasible Feasible