許廠煤礦3.0 Mta新井設計含5張CAD圖.zip,許廠煤礦3.0,Mta新井設計含5張CAD圖,煤礦,3.0,Mta,設計,CAD Control of gas emissions in underground coal mines Klaus Noack* DMT-Gesellschaft für Forschung und Prüfung mbH, Institut für Bewetterung, Klimatisierung und Staubbek?mpfung, Franz-Fischer-Weg 61, Essen, Germany Received 2 August 1996; accepted 24 February 1997. Available online 24 November 1998. Abstract A high level of knowledge is now available in the extremely relevant field of underground gas emissions from coal mines. However, there are still tasks seeking improved solutions, such as prediction of gas emissions, choice of the most suitable panel design, extension of predrainage systems, further optimization of postdrainage systems, options for the control of gas emissions during retreat mining operations, and prevention of gas outbursts. Research results on these most important topics are presented and critically evaluated. Methods to predict gas emissions for disturbed and undisturbed longwall faces are presented. Prediction of the worked seam gas emission and the gas emission from headings are also mentioned but not examined in detail. The ventilation requirements are derived from the prediction results and in combination with gas drainage the best distribution of available air currents is planned. The drainage of the gas from the worked coal seam, also referred to as predrainage, can be performed without application of suction only by over or underworking the seam. But in cases where this simple method is not applicable or not effective enough, inseam-boreholes are needed to which suction is applied for a relatively long time. The reason for this is the low permeability of deep coal seams in Europe. The main influences on the efficiency of the different degasing methods are explained. Conventional gas drainage employing cross measure boreholes is still capable of improvement, in terms of drilling and equipment as well as the geometrical borehole parameters and the operation of the overall system. Improved control of gas emissions at the return end of retreating faces can be achieved by installation of gas drainage systems based on drainage roadways or with long and large diameter boreholes. The back-return method can be operated safely only with great difficulty, if at all. Another method is lean-gas drainage from the goaf. The gas outburst situation in Germany is characterized by events predominantly in the form of ‘non-classical' outbursts categorized as ‘sudden liberation of significant quantities of gas'. Recent research results in this field led to a classification of these phenomena into five categories, for which suitable early detection and prevention measures are mentioned. Author Keywords: gas emission; prediction; pre-degassing; gas drainage; gas outbursts 1. Introduction Coal deposits contain mine gas (mostly methane) in quantities which are functions of the degree of coalification and permeability of the overburden rocks. This is the reason why the gas content of coal seams (and rock layers) varies from 0 m3/t in the flame coal and gas-flame coal of the northwestern Ruhr Basin to >25 m3/t in the anthracite of Ibbenbüren in Germany. When influenced by mining activities this gas is emitted into the coal mine. For better understanding of this process a distinction has been established between basic and additional gas emissions. Basic gas emission is the gas influx from the worked coal seam, which is the equivalent of a partial influx in a multi-seam deposit and of the total gas influx in a single-seam deposit. Additional gas emission represents gas influx coming from neighbouring coal seams (in the case of a multi-seam deposit) and from associated rock layers. The additional gas emission may be in excess of ten times the basic gas emission. So it is mostly the additional gas emission which determines the measures to control the gas emission. In Germany the gas emission is considered to be under control if the gas concentration of the mine air can be kept permanently at all relevant places under 1% CH4. This value is at an adequate distance to the lower explosion limit of methane-air mixtures, which under normal conditions is 4.4% CH4. In exceptional cases, the permissible limit value can be raised to 1.5% CH4. For historical reasons, different permissible limits sometimes apply in other countries, for example 1.25% CH4 in the United Kingdom and up to 2% CH4 in France. Basically, the options for control of gas emission are as follows: (1) Total avoidance of gas release from the deposit. This is only possible with regard to the additional gas emission and only for mining procedures which do not affect stability; hence permeability of the overlying and underlying strata (e.g., room-and-pillar mining where the pillars are left standing during the development phase). (2) Removal of the gas from the deposit before working. For this purpose, all procedures for pre-degassing, either by vertical or by deflected cross measure boreholes drilled from the surface, or by inseam-holes drilled below ground, are technically suitable provided the natural or induced gas permeability permits pre-degassing. (3) Capture and drainage of the gas during mining operations before it mixes with the air flow. This is a classic procedure developed for capturing the additional gas using drainage boreholes, drainage roadways or drainage chambers. (4) Homogenize and evacuate the gas influx after diluting it with sufficient amount of air. This involves panel design, air supply, air distribution, and the prevention of gas outbursts. The following discussions concentrate on problems which are currently given priority in the European Union (EU) funded research. They also cover a significant portion of the gas emission problems worldwide. Problems from non-EU states (e.g., Australia, the Community of Independent States (CIS), South Africa and the United Stated of America (USA)) are also taken into consideration, as far as the author's knowledge permits it. This subject matter is presented in a condensed form under the following headings: prediction of gas emissions; measures taken to control gas emissions; pre-degassing of coal seams; optimization of conventional gas drainage; control of gas emissions for retreating faces; and prevention of gas outbursts. 2. Prediction of gas emissions Prediction of firedamp emission has been practized for many years in the German hardcoal industry (Winter, 1958; Schulz, 1959; Noack, 1970 and Noack, 1971; Flügge, 1971; Koppe, 1975) so that several prediction methods are now available. Among these, the following methods are mentioned: (1) the calculation of the amount of gas emission (Koppe, 1976; Noack, 1985), as used to deal with emission from both the worked coal seam and adjacent seams, which are disturbed by earlier mining activities; (2) the calculation of the reduction of gas pressure (Noack and Janas, 1984; Janas, 1985a and Janas, 1985b), as used in undisturbed parts of the deposit; and (3) prediction methods for the worked coal seam gas emission from longwall faces, for the gas emission from headings and for the gas emission from coal seams cut through during drifting. The first two methods provide a prediction of the specific gas emission from a mine working, expressed in cubic metres of gas per ton of saleable coal production. The gas influx to the mine working in cubic metres of gas per unit time, which is a relevant factor for mine planning, can be derived from multiplying the predicted result by the scheduled production volume. Both methods determine the mean gas emission from a coal face area for a nearly constant face advance rate during a sufficiently long period of time (several months). The prediction assumes that the zone from which the gas is emitted is fully developed, in other words the coal face starting phase has been passed. Furthermore, the coal face has to be above a critical length (i.e., longer than 180–190 m at 600 m working depth and longer than 220–240 m at 1000 m depth). The influx of gas to a coal face area (both into the mine air current and into the gas drainage system) is defined by the following factors: (1) the geometry and size of the zone from which gas is emitted, both in the roof and the floor of the face area, including the number and thickness of gas-bearing strata in that zone; (2) the gas content of the strata; (3) the degree of gas emission, as a function of time- and space-related influences; and (4) the intensity of mining activities. The geometry and size of the zone from which additional gas is emitted are simplified forming a parallelepiped above and below the worked area; its extension normal to the stratification depends on the prediction method. The number and location, type, and thickness of the strata in the zone from which additional gas is emitted can be derived from existing boreholes, staple-shafts, and roadways inclined to the stratification. The gas content of the strata (Paul, 1971; Janas, 1976; Janas and Opahle, 1986) is difficult to determine. There are two alternatives for direct gas content determination available for coal seams (Verlag Glückauf GmbH, 1987). One alternative uses samples of drillings from inseam-boreholes (for developed seams) and the other alternative uses core samples from boreholes inclined to the stratification (for undeveloped seams). Since a suitable method of determining the gas content of rock is not yet available, a double prediction is made with the first prediction neglecting the rock altogether and the second prediction using the assumption of an estimated gas content of the rock strata. The methods for predicting the proportion of gas content emitted are basically divergent. On the one hand the prediction, which is based on the degree of gas emission, assumes that the emitted gas proportion is not a function of the initial gas content but rather of the geometric location of the relevant strata towards the coal face area. The other method, which relies on gas pressure, commences with a fixed residual gas pressure, hence residual gas content. Its value depends on the geometric location of the strata. This means that the emitted proportion of the gas content, representing the balance against the initial gas content, depends on the latter. 2.1. Prediction for previously disturbed conditions The method to predict the total gas make from longwalling in a previously disturbed zone in shallow to moderately inclined deposits (dip between 0 and 40 gon) is based on the degree of gas emission (Fig. 1). It uses the degree of gas emission curve designated as PFG for the roof (considering an attenuation factor of 0.016) and the curve designated as FGK for the floor. Fig. 1. PFG/FGK method. For practical reasons the upper boundary of the zone from which gas is emitted is assumed to be at h=+165 m, whereas, the lower boundary is at h=?59 m. In the absence of empirical data a mean degree of gas emission of 75% in the worked coal seam is assumed. Above the seam, from the h=+0 m level to the h=+20 m level, and below the seam from the h=?0 m level to the h=?11 m level, the degree of gas emission is assumed to be 100%. For the purpose of prediction, the surrounding rock strata are considered as fictitious coal seams for which reduced gas contents are assumed. The reduction factors are 0.019 (for mudstone), 0.058 (for sandy shale) or 0.096 (for sandstone). 2.2. Prediction for previously undisturbed conditions The method to predict the total gas make from longwalling in a previously undisturbed zone is based on the residual gas pressure profiles shown in Fig. 2. There are three zones visible in the roof and two in the floor, which are characterized by varying residual gas pressure gradients. The upper and lower boundaries of the zone from which gas is emitted (hlim and llim, respectively) are defined by the intersection of the residual gas pressure lines and the level of initial gas pressure pu, thus are dependent on the latter. Fig. 2. Gas pressure method: residual gas pressure lines dependent on thickness of the worked coal seam. The breaking points of the residual gas pressure profile for 1 m of worked coal seam thickness (continuous line) are defined by the coordinates in Table 1, whereas the lines are characterized by the residual gas pressure gradients also in Table 1.  Table 1. Parameters for the gas pressure method Full-size table (<1K) View Within Article The dotted line on Fig. 2 applies to 1.5 m of worked coal seam thickness and shows that the h1 and h2 ordinate levels relating to the roof increase in linear proportion to the thickness of the worked coal seam, with gradients declining correspondingly. There is no dependence on coal seam thickness in the floor, where the value of l1 remains constant at ?33 m. Based on the illustrated residual gas pressure profile, the residual gas pressures are first determined layer by layer in accordance with the mean normal distance of a layer from the worked coal seam and afterwards they are converted to residual gas contents using Langmuir's sorption isotherm. The difference between the initial and residual gas contents finally represents the emitted proportion of the adsorbed gas which is the required value. To this value will then be added the free gas, the proportion of which is found by multiplying the effective porosity of the strata under review by its thickness and gas pressure difference. Empirical values have to be used for the effective porosity of coal and rock for methane. Typical values for the coal are between 1 and 10%, and for the rock they are between 0.3 and 1.3%. The values vary in a wide range and depend on chronostratigraphy. In the absence of empirical values for the proportion of gas emission from the worked coal seam a value of 40% would be assumed. 2.3. Comparison of the two methods The gas pressure method may claim the following advantages over the prediction based on the degree of gas emission: There are no rigid delimitations of the upper and lower zones from which gas is emitted. They rather depend on the value of the initial gas pressure and on the type of strata. In the roof the effect of the thickness of the worked coal seam is considered in the profile of residual gas pressure. The prediction takes into account not only the adsorbed gas but also the free gas; this is for both, the coal seams and the surrounding strata. The total gas content rather than the desorbable proportion is used for the prediction. 2.4. Other methods The prediction methods for the worked coal seam gas emission in longwalls and for inseam-headings as well as for coal seam cut through operations during drifting with tunneling machines cannot be explained in detail. For further information refer to the following papers: Noack, 1977; Janas and Stamer, 1987; Noack and Janas, 1988; Noack and Opahle, 1992. It should be mentioned that DMT is testing the prediction of gas emission in machine-driven headings on the base of the INERIS method. Fig. 3 shows an excellent conformity between calculated and measured values (Tauziède et al., 1992). Fig. 3. Comparison between calculated and measured values of gas emission. 煤礦井下瓦斯涌出控制 摘要： 一種先進的方法已在與煤礦井下瓦斯涌出極其相關的領域獲得。雖然如此，卻仍然需要進一步對方法進行改進，如預測瓦斯涌出量，選擇最合適的盤區(qū)設計，延長預抽放瓦斯系統(tǒng)，進一步優(yōu)化生產(chǎn)中抽放瓦斯系統(tǒng)，在在后退式開采時控制瓦斯涌出，以及預防瓦斯突出。介紹和批判性的評價這些重要問題的研究成果。介紹已回采和未回采長壁工作面的瓦斯涌出量預測方法。已開采煤層的瓦斯涌出量和巷道瓦斯涌出量的預測方法也被提及但未深入研究。對通風的要求取決于預測結果，制定基于瓦斯抽放的最佳現(xiàn)有風流分配方案。已開采煤層的瓦斯抽放，同預抽放一樣，不用使用抽風機，而是通過從上面或下面進入煤層來進行。但有時這種簡單的方法不是非常合適或有效，這時將會需要內(nèi)接縫鉆孔，同時會較長時間的使用抽風機，這是由歐洲深煤層的低滲透性決定的。解釋影響不同排瓦斯方法效率的主要因素。傳統(tǒng)的瓦斯抽放技術，無論從鉆眼和設備，以及鉆孔的幾何參數(shù)和操作的整體系統(tǒng)上都有較大的改進空間。通過基于巷道抽放或長而直徑大的鉆孔抽放的瓦斯抽放系統(tǒng)裝置能夠達到改善控制后退式開采工作面末端瓦斯涌出的目的。這種方法能在非常困難的情況下安全實施。另一種方法是從采空區(qū)傾斜式抽出瓦斯。在德國瓦斯突出以“非典型性”突出為主要形式，即突然釋放大量瓦斯。這個領域的最新研究成果將這些現(xiàn)象分為五類，適用于他們的早期檢測和預防措施也被提到。 關鍵詞：瓦斯涌出 預測 預先脫氣 瓦斯抽放 瓦斯突出 1 概述 煤的沉積物包括瓦斯（主要成分為甲烷），其數(shù)量受煤化程度和地表巖層透氣性的影響。這就是在德國為何瓦斯在煤層（以及巖層）的含量，從在西北部Ruhr Basin礦中的長焰煤含量為0m3/t到在Ibbenbüren礦中的無煙煤含量大于25m3/t的原因。 受采礦活動的影響，瓦斯擴散進礦井。為了更好的理解這個過程，需要區(qū)分基本的瓦斯涌出和附加的瓦斯涌出?；就咚褂砍鍪侵竿咚箯拈_采煤層涌出，其等價于在多層沉積物中的部分涌出和在單層沉積物中的全部涌出。附加瓦斯涌出是指相鄰煤層（就多層沉積物而言）和連接巖層的瓦斯涌出。附加瓦斯涌出量可能會超過基本瓦斯涌出量的10倍。因此瓦斯涌出的控制措施主要由附加瓦斯涌出量決定。 在德國，如果礦井空氣中的瓦斯?jié)舛仍谒邢嚓P地段始終處于1%以下，那么則認為瓦斯涌出量在控制之中。這個值離甲烷空氣混合氣體的爆炸下限有足夠的距離，在正常情況下當達到甲烷空氣混合氣體的爆炸下限時甲烷的含量為4.4%，在礦井中甲烷濃度的極限值可允許達到1.5%。由于歷史的原因，其他的國家有時會使用不同的極限值，如在英國礦井中甲烷允許的極限值為1.25%，在德國則為2%。 控制瓦斯涌出的方法基本如下： （1）總體上避免瓦斯從沉積物中釋放。這在僅考慮附加瓦斯涌出或采煤過程不影響穩(wěn)定性時才有可能。 （2）在開采之前從沉積物中除去瓦斯。為了這個目的，所有自然的或人為的預先脫氣的方法，無論是通過通風，還是從地表打斜交叉測量鉆孔，以及從地下打內(nèi)接縫鉆孔，在技術上都是合理的。 （3）煤礦生產(chǎn)期間，在瓦斯與空氣流混合之前對其進行捕捉和抽放。這是為了使用抽放鉆孔，抽放巷道或抽房峒室進行瓦斯捕捉而發(fā)展起來的經(jīng)典方法。 （4）用足量的空氣將瓦斯流稀釋后再將其均質(zhì)和疏散。這涉及到盤區(qū)設計、空氣供應、空氣分配、以及預防瓦斯突出。 接下來的討論將集中于在歐盟提供基金的研究中現(xiàn)在優(yōu)先給予考慮的問題。這其中也包括世界上有關瓦斯涌出問題的重要部分。就筆者許可的范圍內(nèi)，來自非歐盟國家（如澳大利亞、獨聯(lián)體、南非、美國）的問題也將被考慮。本文主題的提出是基于對以下幾個標題的考慮：瓦斯涌出量預測、考出瓦斯涌出的措施、煤層預先脫氣、常規(guī)瓦斯抽放技術的優(yōu)化、后退式開采瓦斯涌出的控制、瓦斯突出的預測。 2 瓦斯涌出預測 甲烷涌出量預測在德國硬煤工業(yè)已經(jīng)實踐了很多年(Winter, 1958; Schulz, 1959; Noack, 1970 and Noack, 1971; Flügge, 1971; Koppe, 1975)。所以現(xiàn)在已經(jīng)得到了一些預測方法。下面將介紹其中的一些： （1）瓦斯涌出量的估算(Koppe, 1976; Noack, 1985)，就像用于處理已采煤層和鄰近煤層的涌出一樣，已被先前的采煤活動干擾。 （2）瓦斯壓力減少量的估算(Noack and Janas, 1984; Janas, 1985a and Janas, 1985b)同樣地被應用于沉積物的未開采部分中。 （3）長壁工作面的已采煤層的瓦斯涌出預測方法，掘進巷道瓦斯涌出預測方法以及在鉆孔時被斜穿的煤層的瓦斯涌出預測方法。 前兩種方法是預測礦井具體瓦斯涌出量的方法，用平均生產(chǎn)一噸煤涌出的瓦斯立方米數(shù)來表示。單位時間內(nèi)瓦斯涌出的立方米數(shù)，可通過預測結果乘以預計產(chǎn)生體積得到，它是制定煤礦計劃的一個重要因素。 這些方法確定了在相當長的時間內(nèi)（幾個月）以幾乎相同的工作面掘進速度工作時采煤工作面的平均瓦斯涌出量。這種預測假定瓦斯涌出的區(qū)域已經(jīng)被全面開拓，也就是說采煤工作面的開始階段已經(jīng)過去。此外，采煤工作面長度必須大于臨界長度（即工作面600m深時，其長度要大于180——190m,1000m深時，其長度要大于220-240m）。 涌進采煤工作面區(qū)域（包括進入井下風流和瓦斯抽放設備）的瓦斯受以下因素的限制：（1）在工作面頂板和底板內(nèi)瓦斯涌出區(qū)域的幾何形狀和尺寸，包括該區(qū)域內(nèi)瓦斯走向?qū)拥暮穸群蛿?shù)量；（2）瓦斯走向?qū)拥耐咚购?；?）瓦斯涌出程度；（4）采煤活動的強度。附加瓦斯涌出區(qū)域的幾何形狀和尺寸被簡化為位于開采工作面上側(cè)和下側(cè)的平行六面體。它距煤層的平均距離取決于預測方法。 附加瓦斯涌出區(qū)域內(nèi)煤層的數(shù)量、位置、種類以及厚度可通過地面鉆孔，喑井-豎井，連接煤層的平巷來得到。煤層瓦斯含量很難確定。對于直接從煤層中獲取瓦斯含量有兩種方法可供選擇（Verlag Glückauf GmbH, 1987）。一種方法利用從內(nèi)接縫鉆孔鉆取得煤樣得到（已開拓煤層），另一種方法利用連接煤層的地面鉆孔內(nèi)的巖心樣品得到（未開拓煤層）。因為沒有一種合適的方法確定巖石中的瓦斯含量，一種雙重預測方法被應用，第一種預測忽略所有巖石，第二種預測應用一種預測巖層瓦斯含量的假想。 預測涌出瓦斯占瓦斯含量比例的方法基本上存在分歧。一方面，基于瓦斯涌出程度的預測方法假定涌出瓦斯量與瓦斯原始含量有關也與接近采煤工作面區(qū)域相關煤層的幾何位置有關。另一種方法，依靠瓦斯壓力，利用瓦斯涌出后的瓦斯壓力得到殘余的瓦斯量。它的結果取決于煤層的幾何位置。這也就是說涌出的瓦斯量占瓦斯含量的比例與原始瓦斯含量無關，而主要取決于后者。 3 已回采條件下的預測 預測薄及中厚傾斜煤層先前開采區(qū)域的長壁工作面中瓦斯總量的方法取決于瓦斯涌出程度(圖1) 。用PFG表示頂板瓦斯涌出程度曲線（考慮0.016的衰減系數(shù)），用FGK表示底板瓦斯涌出曲線。 出于實際的原因，瓦斯涌出區(qū)域的上部邊界被設定為+165米，下部邊界被設定為-59米。在缺乏經(jīng)驗數(shù)據(jù)的情況下，已開采煤層的瓦斯涌出程度被設定為75%。在煤層之上，從高度為0米到+20米，在煤層之下，從高度0米到-11米，瓦斯涌出程度被設定為100%。 為了預測方便，巖層周圍被假想為煤層，并且煤層的瓦斯含量被認為是減少了的。泥巖中的減少系數(shù)為0.019，沙頁巖的減少系數(shù)為0.058，砂巖中的減少系數(shù)為0.096。 1.3未回采條件下的預測 預測未回采區(qū)域長壁工作面瓦斯總量的方法是以殘余瓦斯壓力的基本情況為基礎的（圖2）。由圖可知，在頂板有三個區(qū)域，底板有兩個區(qū)域，他們呈現(xiàn)出不同的殘余瓦斯壓力梯度。 upper boundary 上部邊界 weakened zone 減弱區(qū) cleaved zone 黏著區(qū) caved zone 陷落區(qū) loosened 疏松區(qū) lower boundary 下部邊界 residual gas pressure in bar 殘余瓦斯壓力（bar） 以殘余瓦斯基本情況的圖例為基礎，殘余瓦斯壓力根據(jù)每一層距離開采煤層的平均一般距離一層層確定，然后使用Langmuir等溫吸附式轉(zhuǎn)化為殘余瓦斯含量。由原始瓦斯含量和殘余瓦斯含量的差異最終得到吸附瓦斯的涌出比例，吸附瓦斯量是需要得到的值。然后這個值與游離瓦斯量相加，游離瓦斯量的比例由修正后煤層的有效孔隙率乘以煤層厚度和煤層瓦斯壓力的差得到。經(jīng)驗值用于煤和巖石有效孔隙中的甲烷。典型值在煤中為1%-10%，在巖石中為0.3%-1.3%。這個值在不同的范圍內(nèi)變化。在沒有開采煤層瓦斯涌出比例的經(jīng)驗值時，這個值將被設定為40%。 2.3兩種方法的比較 基于瓦斯涌出程度的瓦斯壓力預測方法有以下優(yōu)點：瓦斯涌出區(qū)域沒有嚴格的上部和下部邊界。他們主要依靠原始瓦斯壓力值和巖層種類。在頂板，殘余瓦斯壓力被認為會影響開采煤層的厚度。不僅預測了吸附瓦斯量，而且還預測了游離瓦斯量；不僅預測了煤層瓦斯量，而且還預測了巖層瓦斯量。在預測中應用到了瓦斯總量，而沒有應用解吸瓦斯比例。 2.4其他方法 長壁工作面已開采煤層瓦斯涌出量以及在隧道掘進機鉆孔期間對內(nèi)接縫-平巷和煤層進行斜穿孔操作時的瓦斯涌出量的預測不能被詳細的解釋。進一步的了解需要參閱以下論文：Noack, 1977; Janas and Stamer, 1987; Noack and Janas, 1988; Noack and Opahle, 1992. 應該提到的是，DMT正在以INERIS方法為基礎對機掘工作面瓦斯涌出量預測方法進行測試。圖3顯示預測值和測量值非常相似。 Measured 測量 calculated 預測 advance 掘進 Volume of CH4 release CH4釋放體積