陳四樓煤礦240萬(wàn)噸新井設(shè)計(jì)【含CAD圖紙+文檔】
陳四樓煤礦240萬(wàn)噸新井設(shè)計(jì)【含CAD圖紙+文檔】,含CAD圖紙+文檔,陳四樓,煤礦,萬(wàn)噸新井,設(shè)計(jì),cad,圖紙,文檔
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綜采工作面過(guò)斷層技術(shù)研究
摘要：斷層是影響綜采工作面煤炭開(kāi)采的重要因素之一。采取有效的技術(shù)措施，高效率地通過(guò)斷層，是提升煤炭采出率、緩解采掘接續(xù)和采裝銜接緊張局面的重要措施，也只有高效地通過(guò)斷層，才能使煤炭開(kāi)采企業(yè)取得更高的經(jīng)濟(jì)效益與社會(huì)效益。
關(guān)鍵詞：綜采工作面;過(guò)斷層;方法措施
0引言
目前，在我國(guó)一次能量消費(fèi)結(jié)構(gòu)中，煤炭占75%以上。煤炭不僅是我國(guó)的基本燃料，又是重要的工業(yè)原料，電力、鋼鐵、石油加工、水泥、化學(xué)原料五大行業(yè)都離不開(kāi)煤炭，因此，煤炭工業(yè)的發(fā)展直接關(guān)系到國(guó)計(jì)民生。為使我國(guó)能源戰(zhàn)略持續(xù)穩(wěn)定的發(fā)展，必須穩(wěn)步高效地發(fā)展煤炭工業(yè)。
我國(guó)是世界上煤炭資源最豐富的國(guó)家之一。據(jù)不完全統(tǒng)計(jì)，己知含煤面積約55000了，探明總儲(chǔ)量在9000億t以上，居世界前列。自1989年，我國(guó)一直是世界第一大煤炭生產(chǎn)國(guó)和消費(fèi)國(guó)，煤炭產(chǎn)量占世界煤炭產(chǎn)量的1/4以上，而淺部煤炭資源已逐漸枯竭，深部資源地質(zhì)條件一般比較復(fù)雜，而目前影響綜采工作面設(shè)備效能發(fā)揮的主要原因之一是地質(zhì)構(gòu)造。在地質(zhì)構(gòu)造(如斷層、褶曲、陷落柱、沖刷帶和火成巖侵入等)中，斷層是最經(jīng)常遇到的，對(duì)生產(chǎn)影響也是最大的一種地質(zhì)構(gòu)造。大落差斷層（）在精確地質(zhì)階段已基本查明，中等落差斷層（）大多數(shù)在采取開(kāi)拓生產(chǎn)準(zhǔn)備階段亦能查明，小落差斷層（）往往不易被發(fā)現(xiàn)，恰恰這些小斷層會(huì)給綜合機(jī)械化工作面回采帶來(lái)不同程度影響，有可能使工作面被迫停產(chǎn)或搬家，或降低工作面單產(chǎn)和月推進(jìn)度或損壞設(shè)備，對(duì)工作面回采帶來(lái)一系列不利影響。而綜采工作面安全過(guò)斷層,可以減少綜采安裝、搬家倒面的次數(shù),避免了時(shí)間和設(shè)備的浪費(fèi)，緩和采掘接替矛盾；回收斷層煤柱，有效提高了礦井煤炭的采出率，降低了煤炭開(kāi)采成本，對(duì)企業(yè)發(fā)展、提高經(jīng)濟(jì)效益起到了促進(jìn)作用。因此有必要對(duì)綜采工作面過(guò)斷層的關(guān)鍵和客觀規(guī)律進(jìn)行理論研究，從而對(duì)綜采工作面過(guò)斷層提出合理的技術(shù)措施。
1從斷層參數(shù)分析斷層對(duì)綜采工作面的影響
綜采工作面過(guò)斷層帶時(shí)，會(huì)對(duì)正規(guī)循環(huán)帶來(lái)較大影響，支護(hù)、移溜、移架的難度將加大，斷層的各種參數(shù)、斷層帶在工作面出露巖石的性質(zhì)及其在工作面空間出露大小決定了這一系列問(wèn)題的復(fù)雜程度。
影響綜采推進(jìn)的斷層參數(shù)包括落差(h)、夾角()和傾角()。
1.1落差大小對(duì)綜采工作面的影響
根據(jù)對(duì)昕益礦綜采工作面調(diào)查，都是10m以下落差的斷層，其中小于1 m的占斷層總數(shù)的12%，1m～2 m的占13%，2 m～3 m的占25%，大于3 m的占50%。
不同大小的斷層落差，對(duì)生產(chǎn)產(chǎn)生的影響程度也不同，當(dāng)落差線性增大時(shí)，工作面的產(chǎn)量呈線性下降。工作面揭露斷層的巖石硬度和裂隙發(fā)育程度等的不同，對(duì)工作面的影響結(jié)果也不相同。當(dāng)工作面揭露巖石穩(wěn)定，用采煤機(jī)落煤時(shí)，對(duì)產(chǎn)量影響較?。划?dāng)工作面揭露松軟巖層，巖層裂隙較多，表現(xiàn)較為破碎，頂板冒落可能增大，從而增加工作面處理冒頂?shù)臅r(shí)間增多，進(jìn)而影響產(chǎn)量；當(dāng)巖石堅(jiān)硬，必須用爆破方法落煤時(shí)，影響60%左右的產(chǎn)量。揭露斷層對(duì)工作面的影響約為推進(jìn)方向6長(zhǎng)。
1.2交角大小對(duì)綜采工作面的影響
據(jù)歷年來(lái)回采面統(tǒng)計(jì)，大部分?jǐn)鄬优c工作面斜交，其中交角100以下的占14%，10°～60°占79%，60°～90°占7%。工作面零交角過(guò)斷層，工作面幾乎在全長(zhǎng)范圍內(nèi)揭露斷層，支架同時(shí)進(jìn)入壓力增大狀態(tài)，如果斷層帶填充物塑性差、揭露斷層上下盤巖石破碎，則通過(guò)斷層危險(xiǎn)性更大，可造成大面積冒頂，壓死支架。當(dāng)工作面與斷層近于垂直時(shí)，斷層在工作面內(nèi)長(zhǎng)時(shí)間存在，工作面支護(hù)條件一直處于不好狀態(tài)，影響支架的支護(hù)、推移，影響采煤機(jī)組的正常通過(guò)，從而很大程度影響產(chǎn)量。
根據(jù)以往經(jīng)驗(yàn)，工作面與斷層交角在0～30°之間時(shí)，回采面有較大幅度產(chǎn)量下降。隨著交角加大，產(chǎn)量下降幅度將變小并趨于穩(wěn)定。被揭露斷層上下盤巖石穩(wěn)定性對(duì)采煤機(jī)落煤有很大影響，當(dāng)交角不變時(shí)，過(guò)破碎巖石斷層工作面產(chǎn)量是過(guò)堅(jiān)硬巖石斷層工作面產(chǎn)量的28%～78%。70°～80°的交角與小交角比較，揭露斷層上下盤巖石的穩(wěn)定性對(duì)產(chǎn)量的影響小。
1.3傾角對(duì)綜采工作面的影響
斷層上下盤接合面的傾角多為50°～85°。這個(gè)區(qū)間的斷層傾角，將使綜采工作面的產(chǎn)量下降約80%，在這樣的傾角下，工作面空間冒落的巖石增加，壓力拱前移，工作面前方煤巖體會(huì)更加破碎，后期開(kāi)采難度增大。其形成原因是在這個(gè)角度范圍內(nèi)，最大切應(yīng)力集中在回采工作面附近，產(chǎn)生裂隙方向同斷層巖石滑動(dòng)面傾向一致，這將造成上覆巖層作用在支架上的壓力減少，同時(shí)作用在前方煤壁上的應(yīng)力增大，將加劇前方煤體的破壞與片幫的發(fā)生，嚴(yán)重時(shí)支架前方有漏頂發(fā)生。
2綜采工作面過(guò)斷層方法
2.1調(diào)整采高法
當(dāng)斷層上下盤頂?shù)装逯g的煤層高度大于設(shè)備允許的最小割煤高度時(shí),采取留頂煤或留底煤的方式降低采高。先估算出工作面距斷層前方所需要的過(guò)渡段,通過(guò)調(diào)整割煤高度來(lái)通過(guò)斷層實(shí)際操作時(shí)采取留頂煤或底煤過(guò)斷層。通過(guò)斷層后逐步見(jiàn)頂見(jiàn)底開(kāi)采,恢復(fù)原采高,采煤機(jī)每刀留頂煤或留底煤的厚度通常為100～150mm。這種方法適合斷層面較小的地質(zhì)條件。
2.2采煤機(jī)強(qiáng)行截割圍巖法
當(dāng)斷層上下盤頂?shù)装逯g的煤層高度小于設(shè)備允許的最小割煤高度時(shí),且工作面頂?shù)装鍑鷰r比較松軟時(shí),采取挑頂或臥底使煤機(jī)截割圍巖強(qiáng)行通過(guò)斷層,要求過(guò)斷層時(shí)工作面的采高大于設(shè)備允許的最小割煤高度,同時(shí)每刀挑頂或臥底量不得超過(guò)150mm。這種方案能夠保證回采工作面推進(jìn)速度,但增大了設(shè)備損壞率,增加了檢修工作量。
2.3放松動(dòng)炮挑頂起底法
當(dāng)斷層上下盤頂?shù)装逯g的煤層高度小于設(shè)備允許的最小割煤高度時(shí),且工作面頂?shù)装鍑鷰r堅(jiān)硬,采煤機(jī)滾筒割不動(dòng)時(shí),必須對(duì)斷層進(jìn)行超前處理,可采取放震動(dòng)炮挑頂或起底的方法。放炮時(shí)要注意炮孔距預(yù)計(jì)的挑頂后工作面頂板的距離,距離太小會(huì)使頂板受到影響,增加工作面通過(guò)時(shí)頂板管理的難度。根據(jù)觀察,深孔爆破裂隙帶的直徑一般為700mm,通過(guò)深孔爆破處理后,采煤機(jī)截割阻力將減小,工作面就可順利通過(guò)斷層。這種方案能夠降低設(shè)備的損壞率,提高了回采率,但工作面安全管理程序復(fù)雜,頂板控制困難。
2.4跳采法
當(dāng)斷層出現(xiàn)在工作面的兩端沿傾斜范圍在30m以下時(shí),在工作面沿煤層正常段與斷層方位平行開(kāi)掘一條全煤巷道與材料巷或運(yùn)輸巷貫通,工作面縮面回采躲過(guò)斷層,隨推采隨接長(zhǎng)工作面的溜子,直至恢復(fù)正常;若斷層出現(xiàn)在工作面的中部,斷層向面內(nèi)延深的全巖不超2m時(shí),用斷層處留垛的跳采方式通過(guò)斷層,然后打通工作面,與兩巷貫通。當(dāng)回采工作面遇到落差較大的傾向斷層時(shí),一般在另一盤重新打切眼,跳過(guò)斷層重新開(kāi)采。此方法打破了工作面的正常循環(huán),且通風(fēng)管理困難,因此應(yīng)用時(shí)需加強(qiáng)過(guò)斷層期間礦壓觀測(cè)，并加強(qiáng)支護(hù)。
3綜采工作面過(guò)斷層的方法選擇
雖然工作面過(guò)斷層的方法相對(duì)較多，但不管采用哪種方法，其選擇的關(guān)鍵均在于工作面與斷層面的夾角和通過(guò)層位的控制。通過(guò)有效地分析，綜采工作面過(guò)斷層應(yīng)考慮的因素主要包括如下幾條：
工作面破碎頂板暴露范圍。工作面與斷層面的夾角應(yīng)盡可能小,但可調(diào)整范圍一般不大。
推進(jìn)前方梁端頂板和煤壁的控制。無(wú)論挑頂起底法或留頂臥底法均應(yīng)保證梁端頂板處于良好支護(hù)狀態(tài),不能造成漏頂事故。
工作面設(shè)備的控制。一般來(lái)說(shuō),根據(jù)斷層上下盤在工作面揭露狀況,一盤若破頂,另一盤可能就會(huì)破底,從而保證工作面刮板輸送機(jī)不會(huì)出現(xiàn)陡坡,減少采煤機(jī)爬坡困難,避免出現(xiàn)支架歪斜和頂梁錯(cuò)茬咬架的現(xiàn)象。
斷層在工作面出口處的控制。回采巷道掘進(jìn)過(guò)程中或經(jīng)后期改造,對(duì)斷層已經(jīng)處理?；夭蓵r(shí),應(yīng)盡量與其處理方法一致,否則會(huì)造成安全出口出現(xiàn)臺(tái)階而不暢通。
4工作面過(guò)斷層的技術(shù)措施
綜采面過(guò)斷層時(shí)，落差、煤層斷失或上或下的方向、工作面與斷層的交角是對(duì)綜采面過(guò)斷層影響較大的因素。所以，在制定安全技術(shù)措施時(shí)，要重點(diǎn)考慮這些因素。
4.1一般技術(shù)措施
落差超過(guò)回采截割高度的2/3的傾斜和斜交斷層，被揭露斷層面出露煤層高度小于液壓支架最小通過(guò)高度的2/3，且被揭露必需硬過(guò)巖層硬度較大，不益采用采煤機(jī)截割、不好打眼放震動(dòng)炮或放震動(dòng)炮后仍不能爆落時(shí)，一般考慮在斷層另一側(cè)開(kāi)切眼，搬家重新安裝或局部開(kāi)巷道繞采；如斷層落差較小，支架能夠通過(guò)；大落差斷層，頂?shù)装鍘r性松軟，或放震動(dòng)炮后易于落矸時(shí)，應(yīng)使工作面與斷層有較大的交角。在圍巖中等穩(wěn)定時(shí)，工作面與斷層合理的交角為20°～30°，圍巖不穩(wěn)定時(shí)，合理的交角為30°～45°。機(jī)組過(guò)斷層應(yīng)采取破頂或壓底，應(yīng)使處理后的底板取同一坡度，易于機(jī)組、支架通過(guò)。對(duì)于落差小于1m的斷層，當(dāng)煤層向上斷失時(shí)，揭露的頂板條件如相對(duì)好些，應(yīng)采取采煤機(jī)與溜子上漂的措施，并留底使采高降低，從而促使支架前梁和頂板盡快在斷層另一盤接頂。當(dāng)頂板破碎易片幫時(shí)，要盡量縮小梁端距，并提前掛雙層金屬網(wǎng)，當(dāng)頂板冒落發(fā)生空頂時(shí)，要插背木板，并要使木板深入頂板底面。對(duì)于落差大于1m的斷層，向上斷失的斷距小于煤厚，縮小采高,多留底煤,使支架上浮，工作面推進(jìn)十幾米距離后,支護(hù)、回采設(shè)備可上升至煤層頂板。支架留底煤上浮時(shí)會(huì)有傾倒事件發(fā)生，應(yīng)采取如下措施：掩護(hù)支架頂板在水平方向上轉(zhuǎn)一角度，使底座后半部分升起一定空隙，伺機(jī)墊入物料，支撐式支架，移架有壓力后，降兩后立柱，使柱腳抬起，伺機(jī)墊入物料，合底座在同一水平面上，保持支架處于穩(wěn)定狀態(tài)，可防止損壞各連接部件及移架彈簧板。
對(duì)于向下斷失落差小于1m的斷層，煤層頂板條件較好時(shí)，支架最小高度大于揭露煤層高度，采取挑頂通過(guò)的辦法。過(guò)斷層前的支架應(yīng)有足夠的伸縮量，方便過(guò)斷層時(shí)降低采高，這就要求慶過(guò)斷層前加大采場(chǎng)高度，相反高度過(guò)大時(shí)支架支不到頂板，采用加支木板或木垛來(lái)支護(hù)上部空間。使支架落后正常位置之后，方便支架能順利下移。煤層頂板破碎時(shí)，用采煤機(jī)掃頂，隨機(jī)掛雙層網(wǎng)，并插木板防止移架前頂板冒漏。如遇向下斷失落差大于1m，小于煤厚的斷層，揭露斷層前加大采場(chǎng)高度，掛雙層網(wǎng)支護(hù)，并在其上插木板，使支架伸到最大支撐高度，作好過(guò)斷層準(zhǔn)備，隨著采煤機(jī)割煤向下過(guò)斷層，支架將回縮降低支護(hù)高度，局部煤層變薄時(shí)，需破頂?shù)装宥^(guò)。
4.2俯采過(guò)斷層的措施
俯采過(guò)斷層，支架和溜子要配合好，下栽角度要一致，即在同一斜面上，不然推移千斤頂活塞桿受力方向不在其中線上，將被推彎，活塞密封圈受擠壓損壞，活塞跳槽難以復(fù)位，此時(shí)可采取降柱法，上提支架底座后沿，使之與溜子在同一斜面一同過(guò)斷層。
俯采過(guò)斷層，工作面要做到煤壁直、溜子直且平，刮板機(jī)下栽角不得超過(guò)16°，避免大量煤炭溜出溜槽，造成無(wú)法裝煤。
4.3仰采過(guò)斷層的措施
在斷層前3m～5m處，每推進(jìn)一個(gè)循環(huán)都要增加滾筒揚(yáng)起的高度，視傾斜角度而定，約10 cm左右。溜子和支架，在移動(dòng)后都要在后側(cè)及時(shí)加墊物料，抬高水平，抑制設(shè)備自重下移，從而加強(qiáng)爬坡能力。使支架縮致最小采高，降低重心，增加穩(wěn)定性，以利于通過(guò)斷層。
綜采面支護(hù)、運(yùn)輸設(shè)備進(jìn)入斷層帶后，溜子逐漸在近水平方向上推進(jìn)，在移架時(shí)可將底座提高至與溜子一個(gè)水平，同時(shí)加大采高，減少留頂煤。
于煤壁松動(dòng)、頂板破碎的斷層，要及時(shí)支護(hù)，梁端要緊貼煤壁，護(hù)幫板及時(shí)伸出交到位，冒空頂處在支架上方用木料棚頂。
仰采過(guò)斷層，工作面在推進(jìn)時(shí)必須保持成直線，絕不能在頂板條件好處超挖，條件不好時(shí)避讓，避免造成工作面應(yīng)力集中，惡化支護(hù)條件?？刂坪门郎仙浇嵌?，應(yīng)在16°以下，避免引起溜子和支架在自重作用下往采空區(qū)側(cè)滑落、傾倒，推溜、移架困難，造成無(wú)法循環(huán)推進(jìn)。
5頂板的控制
綜采工作面在通過(guò)斷層時(shí)，頂板破碎度增加、支架實(shí)際支護(hù)阻力降低、支架產(chǎn)生倒歪、端面發(fā)生冒頂?shù)惹闆r時(shí)有發(fā)生。當(dāng)工作面局部出現(xiàn)落差較大的斷層時(shí)，會(huì)使得整個(gè)工作面難于推進(jìn)。斷層的產(chǎn)狀對(duì)工作面的影響很大。當(dāng)工作面遇到斷層時(shí)，礦壓顯現(xiàn)會(huì)出現(xiàn)異常特征，支架過(guò)斷層破碎帶時(shí)常出現(xiàn)的特征如下：
a. 頂板冒落度明顯增加；
b.支護(hù)阻力降低，但在斷層破碎帶兩側(cè)出現(xiàn)壓力增高現(xiàn)象，表現(xiàn)出壓力拱效應(yīng)；
c.支架工作狀態(tài)惡化。
綜采機(jī)組可采取各種方法通過(guò)斷層，它取決于斷層的落差、煤層的厚度、支架架型大小以及可能被采煤機(jī)截割的軟弱圍巖的總厚度等因素。根據(jù)不同情況可以有三種過(guò)斷層方式:
1） 當(dāng)斷層落差小于煤層厚度與支架最小工作高度之差，采煤機(jī)可直接采過(guò)。如果斷層在工作面推進(jìn)方向向上斷離，可采用仰斜式硬過(guò)的方法。其方法是將頂梁仰抬，采煤機(jī)每割一刀將滾筒提高100～200mm，移輸送機(jī)、拉架時(shí)要抬高機(jī)身和底座，并填入煤、矸及木材。斷層處的頂板破碎時(shí)，要在頂梁上鋪設(shè)金屬網(wǎng)，并采取破碎頂板下的控制頂板措施。如果煤層向下斷離，可采用俯斜式硬過(guò)的方法。其作法是降低采高，留頂煤；采煤機(jī)每進(jìn)一刀下臥100～200mm，墊輸送機(jī)身及支撐底座，使其呈俯角狀。
2） 斷層落差較大時(shí)，需用采煤機(jī)切割頂?shù)装宀拍芡ㄟ^(guò)斷層。若圍巖堅(jiān)硬，則需放震動(dòng)炮或采取爆破挑頂挖底才能通過(guò)。
3） 當(dāng)斷層落差大于采高的1/2或接近煤厚時(shí)，一般不采取硬過(guò)的方式，若采取硬過(guò)的方式會(huì)使得坑木消耗太多，回采效率低下，造成經(jīng)濟(jì)效益下滑，故只能另外開(kāi)掘切眼。
5.1鋼筋網(wǎng)錨噴支護(hù)方式
5.1.1鋼筋網(wǎng)巖巷支護(hù)機(jī)理
鋼筋網(wǎng)巖巷支護(hù)是以鋪設(shè)鋼筋網(wǎng)為主的新的支護(hù)結(jié)構(gòu)，其機(jī)理是通過(guò)錨桿的擠壓作用，形成一定厚度的錨固巖體，鋼筋網(wǎng)的作用是防止圍巖的離層和阻止小塊活石片落，同時(shí)可保證錨桿托盤始終處于鎖緊狀態(tài)，平衡各錨桿受力狀況，保證錨固巖梁的整體性，提高錨固巖梁的抗彎能力，當(dāng)錨固巖體承載受剪切破壞后，它能形成圍包力繼續(xù)承載。經(jīng)鋼筋網(wǎng)組合起來(lái)的圍巖具有支護(hù)三個(gè)方面的要素：一是組合的破碎巖體形成較厚的組合拱；二是破裂圍巖錨固體具有與原巖相近的強(qiáng)度；三是破裂圍巖錨固體具有可縮性。以高壓射流形式噴射的混凝土中的部分灰漿滲入巖體表面縫隙，能使互不聯(lián)系或聯(lián)系較弱的巖塊膠結(jié)成一個(gè)整體，從巖層層面加固圍巖，噴射的混凝土可以密貼在巖石表面將圍巖封閉，防止風(fēng)化和潮解，并且和加固的破碎圍巖形成一個(gè)統(tǒng)一的結(jié)構(gòu)體，互相依存，互相制約，聯(lián)合受力，共同變形。另外，鋼筋網(wǎng)支護(hù)強(qiáng)度較普通金屬網(wǎng)支護(hù)強(qiáng)度大，鋼性好，更
能適宜拱型巖巷造型的需要，利于安裝和服務(wù)年限較長(zhǎng)的優(yōu)點(diǎn)，并且留有容許壓力釋放和留有壓力釋放的時(shí)間和空間，在進(jìn)行鋼筋網(wǎng)支護(hù)完成后待其壓力釋放高峰過(guò)后，巷道變形趨于緩慢時(shí)，再進(jìn)行復(fù)噴，既可補(bǔ)救初噴的破壞部分，還可增厚噴層以阻止或減緩圍巖緩慢持續(xù)的變形。
5.1.2鋼筋網(wǎng)支護(hù)施工工藝
在鋼筋網(wǎng)支護(hù)中使用的鋼筋網(wǎng)長(zhǎng)×寬=2（米）×1.0（米），直徑為5mm鋼筋點(diǎn)焊，網(wǎng)目規(guī)格為40mm×40mm；錨桿：采用45#左旋無(wú)縱筋螺紋錨桿，每套錨桿由桿體、銷釘螺母和托盤主成，錨桿全長(zhǎng)1600mm，錨桿直徑尺寸為20mm，托盤用Q235鋼，厚度不小于6mm，直徑不小于140mm樹(shù)脂錨固劑藥卷必須采用正大公司出廠的產(chǎn)品其中錨固劑藥卷各項(xiàng)指標(biāo)：錨固劑藥卷直徑2325mm，遇有斷層或圍巖破碎使用兩個(gè)藥卷，實(shí)現(xiàn)加長(zhǎng)錨固，錨桿眼底部為速凝藥卷，外部為中速凝固藥卷。藥卷攪拌時(shí)間不得超過(guò)10秒鐘，固化時(shí)間（終凝）不得超過(guò)40秒鐘，錨固劑錨粘膠結(jié)厚度為6mm。
1)工藝流程。
光爆后初噴拱頂→鉆拱部錨桿眼→安裝拱部錨桿→掛鋪拱頂鋼筋網(wǎng)→安裝拱部錨桿→鉆兩幫錨桿眼→安裝兩幫錨桿→掛鋪兩幫鋼筋網(wǎng)→安裝兩幫錨桿→復(fù)噴→監(jiān)測(cè)→噴射混凝土（二次支護(hù)）
2)工藝要求。
（1）初噴：光爆后清理拱部活矸危石，按照中心、腰線檢查巷道斷面規(guī)格后，及時(shí)初噴40mm厚混凝土，然后安設(shè)好單體液壓支柱。
（2）打錨桿眼：打眼前嚴(yán)格按量尺畫(huà)眼布置圖定位，眼位誤差不得超過(guò)100mm。深度應(yīng)與長(zhǎng)度相匹配，深度誤差不超過(guò)50mm，打眼時(shí)應(yīng)在釬子上做好標(biāo)志，嚴(yán)格按錨桿長(zhǎng)度打眼。
（3）安裝錨桿、鋪網(wǎng)。?錨桿眼打到預(yù)定深度后，停止鑿巖機(jī)推力，但給足水30秒撤出釬桿。（幫部眼安裝前應(yīng)用掏勺等工具將眼內(nèi)的巖渣、積水清理干凈）。?用錨桿將樹(shù)脂藥卷送入眼底，使用氣扳機(jī)攪拌藥卷時(shí)間為15~20秒后上錨桿托盤螺母，待固化時(shí)間（7分鐘）后上緊螺冒。?上緊螺母時(shí)，擰緊力矩不小于450N/M，終錨力不小于110KN。鋼筋網(wǎng)搭接長(zhǎng)度不得小于0.2m，并用14#鐵線采用雙絲雙扣隔環(huán)聯(lián)結(jié)。網(wǎng)要緊貼壁面，掛網(wǎng)順序是由拱向兩幫延展，網(wǎng)掛好后撤除臨時(shí)支護(hù)。
（4）復(fù)噴：當(dāng)巷道全斷面完成安裝錨桿、掛網(wǎng)后，立即復(fù)噴60mm厚混凝土。
（5）監(jiān)測(cè)：一次支護(hù)施工后，根據(jù)檢測(cè)儀監(jiān)測(cè)應(yīng)力釋放情況確定二次支護(hù)時(shí)間。
（6）二次支護(hù)：待巷道變形基本穩(wěn)定后，清理破壞了的噴層，然后噴射混凝土至設(shè)計(jì)厚度。
5.2“架噴注錨”支護(hù)方式
先架后錨施工方案：架設(shè)U型棚→初噴壁后充填→打幫部錨桿、鎖腿錨桿→淺孔注漿壁后充填→打頂部錨桿、錨索→復(fù)噴→深孔強(qiáng)化注漿。支護(hù)斷面圖見(jiàn)圖5.1）
5.2.1架設(shè)U型棚
采用超前預(yù)注漿并配合超前撞楔超前管理頂板，防止因放炮震動(dòng)造成掉頂。巷道頂部布置5~7根超前預(yù)注漿錨桿注馬麗散固化松軟破碎圍巖，超前注漿錨桿可兼作撞楔使用，撞楔間距不大于200mm。然后進(jìn)行打眼響炮，找掉浮矸危巖后架設(shè)U36拱形棚，棚距600 mm。采用鋼背板腰幫過(guò)頂。鋼背板采用Φ10mm冷拔鋼筋電弧焊加工而成。每棚搭接處采一個(gè)鐵棚撐，鐵棚撐放在中間一個(gè)卡纜處，鐵棚撐采用寬100 mm、厚10 mm的鋼板加工。
5.2.2初噴壁后充實(shí)
架棚完畢后，初噴混凝土封閉圍巖，并將U型棚后充填密實(shí)，噴厚120 mm，確保將U型棚完整封閉?；炷僚浔龋?黃沙:石子=1:2:2。噴射混凝土作業(yè)緊跟迎頭施工。
5.2.3棚外采用錨桿復(fù)合支護(hù)并用錨桿鎖腿補(bǔ)強(qiáng)
滯后迎頭三棚進(jìn)行幫部錨桿(起拱線以下)支護(hù)，淺孔注漿結(jié)束后一個(gè)圓班進(jìn)行頂板錨桿支護(hù)（淺孔注漿滯后迎頭4 m），錨桿間距為800 mm，排距600 mm。同時(shí)最下一根支護(hù)錨桿兼做鎖腿錨桿使用，配合特制加工的鎖腿卡纜進(jìn)行鎖腿，防止棚腿內(nèi)斂。錨桿規(guī)格：GM22/2800-570，每根錨桿使用三卷Z2550型樹(shù)脂藥卷。
5.2.4淺孔充填注漿
注漿材料采用525#普通硅酸鹽水泥，用于封堵大型裂隙和淺層破碎圍巖體加固。水灰比0.85~1.0，在前探梁之后進(jìn)行注漿。注漿錨桿長(zhǎng)度為1.2 m；注漿壓力一般不超過(guò)1.0 MPa，注漿量以不發(fā)生大量跑漿為準(zhǔn)。注漿順序：從起拱點(diǎn)注漿孔開(kāi)始注漿，盡可能低壓注漿，依次向上把棚后空區(qū)充填滿。
1）采用風(fēng)錘進(jìn)行鉆孔，每排5孔布置，間距1500 mm、排距1 200 mm；鉆頭直徑Φ42 mm，短注漿錨桿孔深1.5 m；
2）注漿錨桿：注漿錨桿長(zhǎng)1.2 m，采用4分鋼管制成，鋼管底端砸成扁狀；前端孔徑Φ=8 mm，后端孔徑Φ=4 mm；采用空心速凝水泥卷封孔，短注漿錨桿封孔深度為0.3 m；
5.2.5關(guān)鍵部位錨索補(bǔ)強(qiáng)
在巷道頂區(qū)及幫部每四棚距（2 400 mm）布置5套錨索（錨索與錨桿錯(cuò)開(kāi)布置）。頂板錨索規(guī)格：Φ17.8×6 300 mm，每排按間距2 000 mm對(duì)稱中頂均勻布置3套；幫部錨索規(guī)格：Φ17.8×4 500 mm，每排兩根，在底板向上800 mm處施工并下扎15°。安裝錨索時(shí)每孔使用一卷K2550和三卷Z2550型樹(shù)脂藥卷，預(yù)緊力120 kN。錨索補(bǔ)強(qiáng)應(yīng)在復(fù)噴及深孔注漿前完成。
5.2.6復(fù)噴混凝土
錨桿、錨索加強(qiáng)支護(hù)安裝結(jié)束后再進(jìn)行一次噴漿(厚度100 mm)，防止錨桿、錨索暴露空氣中銹蝕，并為深孔強(qiáng)化注漿做準(zhǔn)備。
5.2.7深孔強(qiáng)化注漿
1）每斷面7孔布置，排距1 800 mm，間距1600 mm，均勻布置，兩幫最下一根注漿錨桿開(kāi)孔至巷道底不大于300 mm，并向下扎30°；鉆頭直徑Φ=42mm，孔深2.6m；注漿壓力為3MPa。
2）注漿錨桿：長(zhǎng)度2 600 mm，4分鋼管制成，鋼管底端砸扁并擰成麻花狀(長(zhǎng)度200 mm)；鋼管底端1.0 m長(zhǎng)度內(nèi)錯(cuò)開(kāi)鉆孔，孔徑由大逐漸變小，前端孔徑Φ=8 mm，后端孔徑Φ=4 mm；錨桿底部采用一卷Z2550型樹(shù)脂藥卷錨固，并采用空心速凝水泥卷封孔，封孔深度為0.8 m。
3）注漿采用525#普通硅酸鹽水泥，水灰比0.85~1.0，注漿滯后耙矸機(jī)5~10 m實(shí)施。
3.2.8“架噴注錨”支護(hù)特點(diǎn)
“架噴注錨”支護(hù)技術(shù)主要解決了松軟破碎圍巖條件下，無(wú)法采用“先錨后架”或“錨錨(注)”復(fù)合支護(hù)的現(xiàn)實(shí)問(wèn)題，并有以下優(yōu)點(diǎn)：
1）利用棚后噴漿及淺孔注漿充填，解決了松軟破碎圍巖無(wú)法進(jìn)行錨桿支護(hù)甚至復(fù)合支護(hù)的難題；
2）利用先架后錨的施工工藝提高了錨桿的主動(dòng)支護(hù)質(zhì)量及圍巖整體強(qiáng)度，解決了支架與錨桿支護(hù)的不耦合承載問(wèn)題，確保了支架與錨桿同步承載，提高了支護(hù)體系的整體承載能力；
3）錨桿在棚外施工確保了錨桿質(zhì)量同時(shí)有利于監(jiān)督檢查；
4）一次支護(hù)為架棚支護(hù)提高了施工安全環(huán)境，有利于頂板安全；
5）施工工藝靈活，可多工序平行作業(yè)。
6綜采工作面過(guò)斷層控制原則
6.1斷層區(qū)域綜采工作面形態(tài)控制分析
綜采工作面在斷層區(qū)域進(jìn)行回采的速度和效果主要取決于以下4個(gè)方面因素:工作面全巖段長(zhǎng)度和鉆爆施工的難易程度、頂板控制的復(fù)雜程度、工作面設(shè)備的運(yùn)轉(zhuǎn)空間和環(huán)境、工作面和上下兩巷的空間關(guān)系。
1)工作面全巖段長(zhǎng)度和鉆爆施工的難易程度。從理論上講,如果不考慮施工工藝問(wèn)題,工作面按斷層走向剖面的煤層斷裂方向,工作面與斷層走向平行,可以取最短路徑從斷層一盤過(guò)渡到另一盤?？墒侨绻@樣控制工作面形態(tài),那么全工作面都為巖石,即使不考慮其他工藝問(wèn)題,單進(jìn)行全工作面巖石的鉆爆施工,在工藝上也是不可行的。實(shí)際上絕大部分?jǐn)鄬邮┕^(qū)域中(不包括進(jìn)出斷層區(qū)的區(qū)域)工作面同時(shí)暴露斷層的上下兩盤和斷層層面,其伴生的褶曲形態(tài)也同時(shí)在工作面中暴露,并隨工作面推進(jìn)由工作面的一端逐漸向另一端移動(dòng),直至移出工作以外。這樣隨著工作面形態(tài)不同,工作面全巖段的長(zhǎng)度也隨之不同。在綜合考慮其他工藝環(huán)節(jié)和裝備性能后,工作面有一個(gè)最佳曲線,并隨著斷層形態(tài)不同,該曲線的形狀也不同。
2)頂板控制的復(fù)雜程度。為了迅速調(diào)整工作面形態(tài),將工作面擬合到最佳曲線,主要通過(guò)工作面的里偽和外偽形態(tài)及角度。而工作面不論里偽還是外偽都必然影響到支架的形態(tài)。當(dāng)支架的里外偽達(dá)到一定程度時(shí),支架的頂梁支撐點(diǎn)必然前移或后移,使支架由面支撐變?yōu)榫€支撐,如果支架有傾倒趨勢(shì)又會(huì)造成點(diǎn)支撐,支架的工況隨之逐步惡化。而斷層區(qū)無(wú)論是層面還是頂部揭煤區(qū)域,頂板都極為破碎。如果支架工況較差則必然引發(fā)大規(guī)模冒頂,極大影響斷層區(qū)施工速度,所以工作面里外偽的角度控制必須考慮支架的工況問(wèn)題。同時(shí)工作面為快速擬合到最佳曲線,則工作面的角度必然出現(xiàn)劇烈變化,該礦試驗(yàn)過(guò)將工作面角度調(diào)整到25以上,工作面支架出現(xiàn)大面積傾倒,工作面頂板控制狀態(tài)極度惡化。工作面與斷層層面夾角過(guò)小,造成斷層層面在工作面暴露面積過(guò)大,頂板破碎難以控制,經(jīng)常造成大型惡性冒頂。所以在擬合最佳曲線時(shí),必須優(yōu)先考慮支架工況和頂板控制問(wèn)題。同時(shí)在允許的條件下盡量加大工作面和斷層層面間的夾角。
3)工作面設(shè)備的運(yùn)轉(zhuǎn)空間和環(huán)境。除頂板控制論述到的支架工況和運(yùn)轉(zhuǎn)環(huán)境問(wèn)題,同時(shí)采煤機(jī)、刮板輸送機(jī)也存在運(yùn)轉(zhuǎn)空間和環(huán)境問(wèn)題,工作面翻越褶曲,必然進(jìn)行里偽和外偽施工。工作面從正常狀態(tài)開(kāi)始向外偽形態(tài)施工時(shí),如果起坡角過(guò)大,支架和工作面輸送機(jī)間會(huì)出現(xiàn)臺(tái)階高差,嚴(yán)重時(shí)會(huì)造成支架高度合理,而機(jī)道高度不能滿足采煤機(jī)通過(guò)的最小高度要求。所以工作面開(kāi)始調(diào)整形態(tài)時(shí),第一次外偽起坡角不應(yīng)過(guò)大。工作面輸送機(jī)小角度起坡,待支架起坡后,在二次起坡直到工作面外偽角度達(dá)到設(shè)計(jì)要求。
4)工作面和上下兩巷的空間關(guān)系。斷層在與上下兩巷相交時(shí),根據(jù)斷層的產(chǎn)狀不同、與兩巷的夾角不同,在兩巷形成長(zhǎng)度不一的全巖和半巖巷道。工作面的頭、尾在兩巷中經(jīng)過(guò)全巖及半巖巷道施工時(shí),除考慮一般性的施工原則,還要注意以下問(wèn)題:
工作面尾部通過(guò)全巖、半巖巷道,除要注意巖石量最低原則和翻越原則,還應(yīng)考慮到回風(fēng)斷面是否滿足通風(fēng)要求和綜采上出口高度不得低于1.8 m的要求。
一般尾部過(guò)斷層影響巷道時(shí),不宜過(guò)多考慮巖石量問(wèn)題,除保證上出口高度,通風(fēng)斷面外,上端口處頂板支護(hù)不宜出現(xiàn)臺(tái)階,必須考慮到工作面尾部過(guò)斷層層面時(shí)破碎頂板的維護(hù)問(wèn)題。總體來(lái)說(shuō),如果支架支撐高度大于上巷高度,可考慮在支架最大支撐高度滿足情況下,尾部適度下臥;由于工作面輸送機(jī)尾一般位于上巷中,所以通常不考慮工作面尾部抬起高于上巷底板。
6.2控制原則綜述
經(jīng)過(guò)以上理論計(jì)算和機(jī)理分析以及大量的現(xiàn)場(chǎng)實(shí)際工藝現(xiàn)象的統(tǒng)計(jì)分析,得出了指導(dǎo)綜采工作面斷層區(qū)施工、工作面形態(tài)控制的10項(xiàng)原則,現(xiàn)綜述如下:
1)工作面巖石量最低原則。
2)基本上翻越局部穿越原則。
3)破軟不破硬原則。
4)工作面與斷層交角最大化原則。
5)不降低采高原則。
6)上巷保證最小行人空間原則。
7)工作面凸凹角不超過(guò)設(shè)備使用極限原則。
8)下巷保證設(shè)備合理搭接和最小過(guò)貨空間原則。
9)俯仰角保證裝備最小通過(guò)空間原則。
10)整體工藝性配合優(yōu)先考慮頂板控制的原則。
7實(shí)例分析
4303工作面是長(zhǎng)平礦首個(gè)大采高工作面,當(dāng)推進(jìn)至機(jī)頭694m,機(jī)尾700m處,在60#支架處揭露SF39(340<50,H=2 5m)正斷層,為保證4303工作面能順利通過(guò)SF39斷層,在詳細(xì)分析該處地質(zhì)資料的基礎(chǔ)上,提出了一套行之有效的措施,為大采高工作面順利通過(guò)斷層提供了技術(shù)保障。
7.1斷層處地質(zhì)概況
4303大采高工作面沿底板推進(jìn),走向長(zhǎng)1030m,傾斜長(zhǎng)225m,煤層平均厚度為5 86m,煤層傾角5~16,平均傾角10;直接頂為砂質(zhì)泥巖,厚度4 85m,局部有裂隙,松軟,含大量白云母片和植物化石。
當(dāng)工作面推進(jìn)至機(jī)頭694m,機(jī)尾700m處,在60#支架處揭露SF39正斷層,如圖1,圖2所示。其對(duì)工作面的影響范圍為25#~130#支架,對(duì)推進(jìn)方向的影響范圍約為42m。工作面采用SL500雙滾筒采煤機(jī)、SGZ1000/1710刮板機(jī)、SZZ1200/375轉(zhuǎn)載機(jī)、DSJ1400/250/3 400膠帶機(jī)、PCM 375型錘式破碎機(jī),從工作面機(jī)頭到機(jī)尾分別布置ZYT 12000/28/62型排頭架4架,ZYGT 12000/28/62型過(guò)渡架2架,ZY12000/28/62型中間架119架,ZYGT12000/28/62型過(guò)渡架2架,ZYT12000/28/62型排尾架3架,共計(jì)130架。
7.2工作面過(guò)斷層技術(shù)措施
7.2.1支架抬起點(diǎn)、抬起坡度、最小采高確定
為了實(shí)現(xiàn)大采高工作面快速、安全地通過(guò)斷層,必須消除或減弱割矸的影響。當(dāng)工作面從斷層的上盤向下盤推進(jìn)時(shí),不是到了斷層面后再向上抬起推進(jìn),而是當(dāng)工作面距斷層面一定距離時(shí)開(kāi)始以一定的仰角推進(jìn),根據(jù)煤層頂?shù)装宓膸r性割頂板巖層或頂煤,留底煤推進(jìn)回采,使得工作面推進(jìn)到斷層下盤時(shí),其底板正好為斷層下盤底板或割矸厚度在1m以下,從而很快恢復(fù)到正常回采,消除或減小了斷層割矸對(duì)工作面回采的影響。因此,支架的起坡點(diǎn)、抬起坡度和最小采高的確定十分重要。
1） 工作面支架起坡點(diǎn)確定
如圖7.3所示,工作面起坡點(diǎn)位置確定采用如下計(jì)算方法:
式中：h——斷層斷距
M——為工作面采高
——為支起坡角
為斷層傾角
2） 工作面支架抬起坡度確定
大采高工作面由于采高增大支架穩(wěn)定性較差,若過(guò)斷層時(shí)支架起坡角太大,梁端距變大,伸出伸縮梁也不能滿足支護(hù)要求,易發(fā)生端面冒落;另外,支架仰角太大易導(dǎo)致支架立柱向采空區(qū)側(cè)傾角變大,形成倒傘形,支架升起越高,支架前梁越下傾,前梁下方采高變低,甚至機(jī)組無(wú)法通過(guò),如圖4所示。根據(jù)支架性能和實(shí)際地質(zhì)情況,大采高工作面仰(俯)角最后確定不超過(guò)7為易,即每刀煤抬高控制在100mm,梁端距最大增至875mm,把支架伸縮梁伸出即可滿足安全生產(chǎn)需要。
4303工作面在推進(jìn)過(guò)程中,早班工作面開(kāi)始抬起,工作面割底矸厚度大約為2 7m,4d后工作面割底矸厚度大約為0 5m。工作面共割煤7 5刀,實(shí)際上每刀煤抬高了0 29m。實(shí)際測(cè)定III4303過(guò)構(gòu)造各支架角度,工作面36#~76#的支架平均仰角超過(guò)了18,最大仰角超過(guò)了21。最終造成65#~75#支架后傾嚴(yán)重,支架前梁下傾嚴(yán)重,下方采高不足3 9m,支架推移困難,機(jī)組無(wú)法正常通過(guò)端面冒頂。最后采用打貼幫柱把支架升起,才逐架勉強(qiáng)通過(guò)。所以設(shè)計(jì)每個(gè)循環(huán)抬起100mm坡度是合理的。
3） 工作面最小采高確定
過(guò)斷層的難易程度與采高的關(guān)系較大,采高越小工作面片幫越少,礦壓顯現(xiàn)越緩和,過(guò)斷層越容易。工作面過(guò)斷層時(shí)的最小采高用下式確定:
式中,為采煤機(jī)滾筒直徑、采煤機(jī)機(jī)身高度,mm;為支架頂梁高度,mm;為護(hù)幫板收回高度、支架最小高度,mm。200mm為空隙余量,其中取二者的最大值。
4303工作面SL500機(jī)組機(jī)身高度為2 6m,采煤機(jī)滾筒直徑為3 2m,支架最小高度為2 8m,頂梁高度為350mm,(一二級(jí))護(hù)幫收回高度為350mm,另外,考慮頂板下沉等不確定因素,最終確定本工作面設(shè)備允許通過(guò)的最小煤層高度為4.5m。
7.2.2 安全技術(shù)措施
SF39斷層面與工作面交線(簡(jiǎn)稱斷層線)至4206巷段為上盤,斷層線至4205巷段為下盤。揭露斷層時(shí),22#~60#支架割矸為斷層上下盤過(guò)渡段。為了順利通過(guò)斷層,25#支架至斷層線段逐步抬高,沿工作面底板(現(xiàn)斷層下盤底板)正?；夭?斷層線至機(jī)尾必須從斷層的上盤逐步抬高,進(jìn)入斷層下盤,沿工作面底板正?；夭?。工作面在往前推進(jìn)時(shí),斷層線逐步向后推移,當(dāng)工作面向前推進(jìn)42m時(shí),在4206巷消失,斷層過(guò)完。具體措施如下:
1)工作面從割矸處開(kāi)始抬起底板,每刀煤抬升高度不得超過(guò)100mm,抬升過(guò)程中要保持平穩(wěn)過(guò)渡。25#支架至斷層線逐步抬起沿工作面底板(現(xiàn)斷層下盤底板)正?；夭?斷層線至機(jī)尾逐步抬起,留底煤推進(jìn)。
2)工作面構(gòu)造區(qū)采高降低至4 5m左右,前后鄰架要平穩(wěn)過(guò)渡,相鄰架間側(cè)護(hù)板錯(cuò)差以不超過(guò)側(cè)護(hù)板本身的2/3為原則,最大錯(cuò)差不得超過(guò)全部側(cè)護(hù)板高度。
3)斷層破碎帶提前采用注漿煤體固化和錨索加固技術(shù),即在工作面漏矸和斷層帶的煤巖體中每隔3~5m打孔,高壓注入諾米加固一號(hào),粘結(jié)固化破碎巖體使之成為整體,然后在打錨索進(jìn)行整體加固,使之成為一個(gè)整體,提高了圍巖的完整性。在注漿過(guò)程中,要注意觀測(cè)孔周圍煤壁情況,隨著反應(yīng)進(jìn)行,壓力逐漸增高,待壓力增高7M P a以上,即可結(jié)束。推進(jìn)度與注漿加強(qiáng)厚度要精確計(jì)算,保證注漿加強(qiáng)體不被割破,每走6個(gè)循環(huán),加固一次破碎帶,對(duì)其他頂板和煤壁片幫嚴(yán)重的也要注漿加固。
4)大采高工作面采高超過(guò)4 5m以上,發(fā)生端面冒頂時(shí),采用普通綜采工作面冒頂時(shí)采用的勾頂、打貼幫柱、伸縮梁上挑板梁等措施,已經(jīng)不能正確處理大采高工作面冒頂。根據(jù)現(xiàn)場(chǎng)的實(shí)際情況采用了單體柱一端頂在支架一級(jí)護(hù)幫板上,另一端頂在支架伸縮梁的尾端(以下簡(jiǎn)稱一級(jí)護(hù)幫加強(qiáng)柱),用單體柱增加支架護(hù)幫板支護(hù)強(qiáng)度,來(lái)封住漏矸口。拉架時(shí),二級(jí)護(hù)幫板下打戧柱,回掉一級(jí)護(hù)幫板的加強(qiáng)柱,開(kāi)始拉架,拉架結(jié)束恢復(fù)一級(jí)護(hù)幫加強(qiáng)柱,最后回收二級(jí)護(hù)幫下戧柱,通過(guò)機(jī)組,逐架通過(guò),直至結(jié)束。此措施一級(jí)護(hù)幫加強(qiáng)柱捆綁危險(xiǎn)系數(shù)高,在制定安全技術(shù)措施時(shí)必須專項(xiàng)說(shuō)明。
5)當(dāng)采高降到4 5m時(shí),采用挖柱窩施工貼幫柱的方式幫板,保證機(jī)組通過(guò)時(shí),不冒落矸石。
6)工作面未能及時(shí)抬起,割矸厚度變厚時(shí),放震動(dòng)炮割矸推進(jìn),具體措施很常見(jiàn),這里不再贅敘。
7)合理調(diào)整工作面機(jī)頭機(jī)尾推進(jìn)度,調(diào)整工作面支架直線度和擺向情況下,保證工作面三直兩平,不咬架、工作面刮板輸送機(jī)不竄前竄后,支架推移不受影響。多推進(jìn)機(jī)頭,少走機(jī)尾,工作面與斷層面大角度相交,防止斷層面在工作面大面積同時(shí)揭露,把破碎帶影響范圍減至在回采控制范圍以內(nèi)。
8)必須保證工作面支架的初撐力達(dá)到合格值,充分利用支架來(lái)支撐頂板。
參考文獻(xiàn)
［1］李平．綜采工作面過(guò)斷層采煤技術(shù)實(shí)踐［J］．煤礦開(kāi)采，2001年3月．
［2］郭守泉，彭永偉．綜采工作面過(guò)斷層技術(shù)綜述［J］．煤礦開(kāi)采，2008年8月．
［3］勾攀峰，胡有光．?dāng)鄬痈浇夭上锏理敯鍘r層運(yùn)動(dòng)特征研究［J］．采礦與安全工程學(xué)報(bào)，2006，23（3）：285－288．
[4]劉建國(guó).大采高單體液壓支柱工作面支護(hù)型式實(shí)踐[J].煤炭技術(shù)，2009，（7）：162-163.
[5]劉鴻博，鄧慶鵬.綜放工作面過(guò)斷層開(kāi)采技術(shù)[J].煤炭技術(shù)，2008，（1）：48-49.
[6]吳元峰，高占峰.一個(gè)工作面兩種回采工藝同時(shí)生產(chǎn)過(guò)斷層方法[J].煤炭工程，2007，（11）：54-55.
[7]張效春.綜采工作面過(guò)斷層、過(guò)沖刷技術(shù)及有關(guān)參數(shù)的確定[J].山西煤炭，2005，25（4）:33-35.
[8]宋延力，魏巍.回采工作面過(guò)斷層的處理方法[J].山東煤炭科技，2005（4）:6-7.
[9]樊繼強(qiáng)，趙軍.綜采工作面過(guò)斷層開(kāi)采技術(shù)探討[J].山東煤炭科技，2002（1）:47-49.
[10]劉錦榮，何富連.大采高綜采工作面支架一圍巖系統(tǒng)穩(wěn)定性探討.煤礦開(kāi)采，1995 (3)，36-40
[11]姜福興，王同旭，潘立友等。礦山壓力與巖層控制.北京:煤炭工業(yè)出版社，2004
任務(wù)書(shū)
學(xué)院 礦業(yè)工程學(xué)院 專業(yè)年級(jí) 學(xué)生姓名
任務(wù)下達(dá)日期：20XX年1月8日
畢業(yè)設(shè)計(jì)日期：20XX年3月12日 至 20XX年6月8日
畢業(yè)設(shè)計(jì)題目：陳四樓煤礦240萬(wàn)噸新井設(shè)計(jì)
畢業(yè)設(shè)計(jì)專題題目： 綜采工作面過(guò)斷層技術(shù)研究
畢業(yè)設(shè)計(jì)主要內(nèi)容和要求：
以實(shí)習(xí)礦井陳四樓煤礦條件為基礎(chǔ)，完成陳四樓煤礦2.4Mt/a新井設(shè)計(jì)。主要內(nèi)容包括：礦井概況、礦井工作制度及設(shè)計(jì)生產(chǎn)能力、井田開(kāi)拓、首采區(qū)設(shè)計(jì)、采煤方法、礦井通風(fēng)系統(tǒng)、礦井運(yùn)輸提升等。
結(jié)合煤礦生產(chǎn)前沿及礦井設(shè)計(jì)情況，撰寫(xiě)一篇關(guān)于綜采工作面過(guò)斷層技術(shù)研究的專題論文。
完成2010年《國(guó)際巖石力學(xué)與采礦科學(xué)雜志》上與采礦有關(guān)的科技論文翻譯一篇，題目為“Hydraulic fracturing after water pressure control blasting for increased fracturing”。
院長(zhǎng)簽字： 指導(dǎo)教師簽字：
翻
譯 部
分
英文原文
Hydraulic fracturing after water pressure control blasting for
increased fracturing
Bingxiang Huang,Changyou Liu,Junhui Fu,Hui Guan
School of Mines,China University of Mining and Technology,South 3rd Ring Road,Xuzhou,Jiangsu 221116,China
Abstract:Traditional hydraulic fracturing techniques generally form main hydraulic cracks and airfoil branch fissures,but main hydraulic cracks are relatively few in number.Hydraulic fracturing after water pressure control blasting can transform the structure of coal and rock mass.Experiments prove that it is an effective method for increasing the number and range of hydraulic cracks,as well as for improving the permeability of coal seams.The technical principle is as follows.First,a hole is drilled in the coal seam and is injected with a gel explosive(a mining water-proof explosive).Then,water is injected into the hole to seal it,at low enough pressure to prevent cracks from forming.Third,water pressure blasting is done by detonating the explosive.The water shock waves and bubble pulsations produced by the explosion cause a high strain rate in the rock wall surrounding the hole.When the stress imposed on the rock wall surrounding the hole exceeds its dynamic critical fracture strength,the surrounding rock breaks and numerous circumferential and radial fractures propagate outward.Lastly,water injection processes,such as general injection,pulse injection,and/or cyclic injection,are carried out to promote hydraulic fracturing.Depending on the fissure water pressure,detonation fissures continue to expand and additional hydraulic fractures with a wider range are formed.Under the effect of detonation pressure,joints and fissures in the coal mass open and propagate,leading to reduced adhesive forces on structural surfaces and thereby enhancing coal cutting.Therefore,this method improves the permeability of the coal seam,effectively weakens the strength of the coal and rock mass,and reduces the surrounding rock stress of the weakened area,effectively solving the problem of having a small number of big cracks.It is a useful technical approach for improving top coal caving,preventing rock burst,preventing coal and gas outbursts,and raising the gas extraction efficiency in colliery.
Key words:Hydraulic fracturing;Water pressure blasting;Crack propagation
1 Introduction
Low-permeability coal-seam gas extraction;hard,thick coal-seam fully mechanized top coal caving;and rock burst control are technical challenges in colliery at present.Hydraulic fracturing is an effective technical approach to resolve these challenges[1].The structure of coal and rock mass is altered through hydraulic fracturing,which can increase cracks in coal and rock mass improve permeability,and weaken strength to reduce any rock bursting liability.After decades of development,more study on hydraulic fracturing has been conducted both in China and else-where[2–14].Simulation experiments and field investigations of hydraulic fracturing show that the traditional hydraulic fracturingin number.In the case of homogeneous rock,a single hydraulic main crack is generally generated and cracks are mainly concen-trated in a band around the hydraulic main crack,whose extent is
small.However,to improve hard,thick top coal cavability,handle hard roof,prevent rock burst,increase permeability of gassy coal seams,and prevent coal and gas outbursts,full re-formation of the structure of coal and rock mass by hydraulic fracturing is needed.This requires that hydraulic fracturing produce more hydraulic cracks,i.e.,increase the number of hydraulic cracks.Therefore,there is an urgent need to study hydraulic control fracturing technology to increase the number of hydraulic cracks,which has important theoretical and practical significance in guaranteeing efficient and safe colliery production.Common explosives blasting for gassy coal seams has safety risks,so they are not suitable.Water pressure blasting,developed
in the past century as a kind of controlled blasting method,can effectively control the generation of blasting flying rocks,air shock waves,blasting tremors,and detonation toxic gases[15–18].Water pressure blasting is a gun-hole blasting technology that uses water as a coupling medium between the cartridge and the charge hole to techniques mainly form water pressure main cracks and airfoil branch fissures,but water pressure main cracks are relatively fewin number.In the case of homogeneous rock,a single hydraulic main crack is generally generated and cracks are mainly concern-trated in a band around the hydraulic main crack,whose extent is small.However,to improve hard,thick top coal cavability,handle hard roof,prevent rock burst,increase permeability of gassy coal seams,and prevent coal and gas outbursts,full re-formation of the structure of coal and rock mass by hydraulic fracturing is needed.This requires that hydraulic fracturing produce more hydraulic cracks,i.e.,increase the number of hydraulic cracks.Therefore,there is an urgent need to study hydraulic control fracturing technology to increase the number of hydraulic cracks,which has important theoretical and practical significance in guaranteeing efficient and safe colliery production.
Common explosives blasting for gassy coal seams has safety risks,so they are not suitable.Water pressure blasting,developed in the past century as a kind of controlled blasting method,can effectively control the generation of blasting flying rocks,air shock waves,blasting tremors,and detonation toxic gases[15–18].Water pressure blasting is a gun-hole blasting technology that uses water as a coupling medium between the cartridge and the charge hole to transfer the explosion pressure and energy at the moment of the explosion to break up rock.The principal characteristics of water are exploited as follows.Since water is difficult to compress,
deformation energy losses are low and energy transmission efficiency becomes high.Water acts to deliver uniform pressure, making the pressure on the surrounding medium relatively smooth and evenly distributed,leading to even breaking of the surrounding rock and greatly reducing the harmful effects of blasting.However,the compression ratio of water exceeds that of rock under high pressure and water also acts as the buffer layer between the explosive products and the rock mass.Not only does this buffer layer extend the interaction time of the shock wave on the rock,
but it also can reduce or eliminate the energy loss in the plastic deformation zone generated in the rock mass.Water pressure blasting is currently a more mature technology in fields such as tunnel excavation and project demolition.In recent years,the application of water pressure blasting to colliery has started in China and elsewhere[19,20].In the former Soviet Union,coal-seam pre-injection internal explosions were conducted by using an 8-m-deep hole of 40 mm diameter to prevent coal and gas outbursts in a gently inclined thin coal seam and a medium thick coal seam.In China attempts were made to create cracks by water pressure blasting to improve the coal-seam gas drainage rate[21].
In view of the problems of existing technology,a preliminary test has been conducted to exploit the advantages of water pressure blasting and hydraulic fracturing.The test results show that hydraulic fracturing after water pressure blasting can increase the number and range of hydraulic cracks efficiently. Based on preliminary studies and test results,the author has proposed the use of water pressure control blasting for increasing permeability and weakening strength as a result of hydraulic fracturing.
2.Using water pressure control blasting to increase permeability through hydraulic fracturing
Water pressure control blasting induces hydraulic fracturing in the borehole of a coal-rock seam,which changes the structure of the coal-rock mass and increases the number and range of hydraulic cracks,thereby increasing permeability and weakening strength.The technique involves the following steps:
(a)Drill a borehole for hydraulic fracturing weakening with a drilling rig,inject an adequate amount of gel explosive(a water-proof mine explosive),and pull the lead wire out of the borehole.
(b)After sealing up the borehole orifice with hole packer or cement mortar,inject water into the hole until it fills the hole or reaches a pressure value below that which would generate water pressure cracks.At this moment,the initial water pressure in the borehole must be less than the orifice rupture water pressure:
where is the minimum principal stress of the crustal stress field around the borehole, is the maximum one,and is the tensile strength of the borehole rock.
(c)Detonate the explosive to carry out water pressure blasting.The water shock waves and bubble pulsations produced by the explosion will cause a high strain rate in the rock wall surround-ing the hole.When the stress imposed on the surrounding rock wall exceeds its dynamic critical fracture strength,the rock ruptures and generates abundant circumferential and radial fractures surrounding the borehole.Meanwhile,because of the rock’s elasticity,the hole’s influence on the surrounding rockstress distribution is about 3–5 times the borehole diameter.Under the effect of subsequent water pressure,cracks are initiated in the wall of the hole when the effective tangential tensile stress of the wall exceeds the rock tensile strength. However,for a given crustal stress field,the position of the maximum effective tangential tensile stress of the borehole wall is a constant.Therefore,to increase the difference of hydraulic crack initiation between the follow-up borehole hall and the blasting cracks and make the blasting cracks craze preferentially,the length of blasting cracks must be greater than 3–5 times the borehole diameter.
(d)Then,perform water injection processes such as general injection,pulse injection,and cycle injection to carry out hydraulic fracturing.Depending on the fissure water pressure,blasting cracks continue to expand and more water pressure fractures with a wider range are formed.
The surrounding rock loosing zone of colliery roadway or grotto for constructing a borehole is generally 1.5–2.0 m.Because the water pressure induced by water pressure blasting is great,the sealing length in the complete surrounding rock section of the borehole must be greater than 2 m.The borehole length for installing the gel explosive must exceed 1 m.Thus,the underground fracturing borehole depth in colliery should not be less than 5 m.
The structure of coal and rock mass is re-formed by hydraulic blasting control fracturing,leading to an increased number of hydraulic cracks,an increase in the permeability of the coal seam, an efficient weakening of the strength of coal and rock mass,and a reduction in the surrounding rock stress of the weakened area.This effectively solves the problem of having a small number of big cracks.There are a number of beneficial effects from this process.The weakening of the hard coal can improve top coal capability,reduce the risk of rock burst,increase the range of coal-
seam fracturing cracks,make gas extraction easier,and prevent coal and gas outbursts,all of which are important in guaranteeing efficient and safe colliery production.
3.Experimental scheme
3.1.Experimental system
We developed a true triaxial hydraulic fracturing experimental system.The system consists of an experiment-bench framework,a loading system,and a monitoring system.The main technical indicators are as follows:(1)the true triaxial stress is loaded on cubic samples to simulate crustal stress;the pressure from the loading plate in three directions can reach 4000 kN.(2)The size of the cubic specimen is or .(3)The water pressure for borehole fracturing can reach 70 MPa.
During borehole fracturing,parameters such as water(liquid) pressure and flow are monitored by an Intelligent Vortex Flow-meter connected to the computer,using established procedures for data collection and storage.During the fracturing simulation, the crack propagation process and geometric morphology are monitored by a Disp-type 24-channel acoustic emission instrument,an RSM acoustic instrument,and a TDS-6 Micro-seismic acquisition system.
3.2.Experimental method
The simulation experiment adopts a side length of 500 mm for the cubic specimen mixed with coal and briquette.(The original coal size is about .)A parameter test of the mechanical properties of both coal and briquette of different ratios has been conducted to ensure that the stiffness,strength,and other properties of the briquette are as similar to coal’s as far as possible.The quality ratio of the simulated sample is determined as coal powder:cement:plaster:water?0.5:1:1:0.8 and its mechanical properties are shown in Table 1.After the specimen has naturally dried,a borehole of 30 cm in length is drilled at the center of the upper surface of the specimen and then SHZ bar glue is used to bond the device bond to the borehole wall to complete the sealing while the sealing depth reaches 20 cm.
We originally planned to use electric detonators to carry out the simulation experiment of hydraulic blasting control fracturing.The explosive amount(1 g)in each electric detonator is modest and the detonators can be detonated in water to achieve the purpose of blasting after sealing under water pressure.Thus, an electric detonator is the ideal blasting equipment for the simulation experiment.However,because the public security sector strictly controls electric detonators,it is hard to obtain blasting electric detonators.Therefore,the large firecracker shown in Fig.1b was used as blasting equipment for the simulation experiment.The hydraulic blasting control fracturing is simplified into two stages to simulate(1)blasting in the borehole and(2)hydraulic fracturing.
The simulated stress field condition is ,,and and the stress direction is shown in Fig.1c.Red poster dye is added to the water tank to make it easier to observe the hydraulic fracture morphology. During the experiment,a microseismic instrument is used to monitor microseismic information of the specimen;the trigger threshold(STA/LTA ratio)of a microseismic event is 1.2,and the amplitude range reaches 500 mA with an STA/LTA time window of (0.1 s)/(1 s).At the same time,acoustic emission and electromagnetic radiation are monitored during the experiment.An acoustic emission probe(R.45)placed in the experimental framework closely sticks to the specimen and an electromagnetic radiation probe stays close to the outer steel ring of the experimental framework.An acoustic emission instrument uses the acoustic emission probe and the electromagnetic radiation probe to take samples at the same time.The frequency domain f of the electro-magnetic radiation probe is 30 kHz.The sampling frequency both the pre-amplifier(BP-SYS)and the acoustic emission probe is 5 MHz;the trigger threshold of the electromagnetic radiation probe is 20 dB,the trigger threshold of the acoustic emission probe is 39 dB,and the pre-amp gain is 60 dB for both.The high-pass filter of the electromagnetic radiation probe is set to 20 kHz and the high-pass filter of the acoustic emission probe is 1 kHz.The low-pass filters for both are set to 400 kHz.
To compare with the results of common hydraulic fracturing in coal and rock mass,one common hydraulic fracturing simulation experiment of fissured coal and rock mass under the same simulated crustal stress and quality ratio of sample has been conducted.
4.Analysis of results
4.1.Crack propagation process of hydraulic fracturing after water pressure control blasting
4.1.1.Blasting
After the large firecracker shown in Fig.1b is lit,it is put at the bottom of the drillhole and then the square iron pad of 70.2 kg containing the experiment framework is set to cover the orifice area of the specimen.The typical microquakes monitored during the experiment are shown in Fig.2,where the abscissa plots time, every small division stands for 0.1 s,and the total time shown is 5 s.In the figure,the first event is the quake caused by the square iron pad after lighting the firecracker;the second event is the microquake event caused by blasting;the third event marks the upward jump of the iron pad caused by the detonation gas after the explosion.
After six blasts at the bottom of the drillhole,the specimen surface shows no visible cracks and is still integrated.The specimen is then placed on the test desk for the hydraulic fracturing
experiment after sealing.
4.1.2.Hydraulic fracturing
The water pressure and acoustic–electric effect during hydraulic fracturing after blasting are shown in Fig.4.A total of seven water injection fracturing experiments were conducted.For the first two,the pressure was controlled manually;for the last five,a high hydraulic pressure was pre-set by a stabilizer and water injection fracturing with high flow was carried out by pressure output switches.When the hydraulic pressure of the first water injection fracturing reaches 1.1775 MPa,a turning point in the hydraulic pressure curve appears(Fig.3b).At this moment,both the pulse number and the amplitude of the electromagnetic radiation show a small peak(Fig.3e and f),indicating that the drillhole wall ruptures(or the original blasting cracks open and burst),meaning that the hydraulic pressure of rupture is 1.1775 MPa.After the hydraulic pressure reaches 1.4775 MPa,it then decreases,showing that the hydraulic fracture propagates at this time.After the hydraulic pressure reaches a maximum of 1.5375 MPa,it falls to 1.4075 MPa with a relatively high speed,meaning that the hydraulic fracture propagates with a large scale.
When the hydraulic pressure becomes about 1.40775 MPa,it remains constant for 9 s and then sharply declines.Meanwhile,both the pulse number and the amplitude of electromagnetic radiation have significant peaks and the deformation and failure of coal and rock mass are exacerbated.In the second water injection by manual control,when the hydraulic pressure reaches 1.2575 MPa,the same situation as with the first injection fracturing appears.The hydraulic pressure exhibits a turning point,which indicates renewed opening of the hydraulic crack.Afterward,the hydraulic pressure rises to 1.4075 MPa and it declines stably in only 3 s.The hydraulic crack perforates through the specimen surface fully;water comes out of the specimen orifice surface(upper surface)and the hydraulic pressure decreases sharply.
During the subsequent fracturing of multiple injections,the ratio of water filtration decreases relatively because of high flow.So the hydraulic pressure reaches a maximum of 1.6775 MPa,which is greater than the maximum pressure obtained by manual control.Thus,when the filtration rate of the coal and rock seam is large,a high flow of water injection fracturing should be used to ensure higher water pressure on the crack tip to cause the hydraulic fracture to propagate.
During the whole process of hydraulic fracturing,seven microquake events were monitored;a typical example is shown in Fig.4.In comparison to blasting quakes,microquakes induced by hydraulic fracture propagation are much weaker.Under laboratory conditions,because the layout space of the probes is small(2 m or less),the difference in time at which each probe receives the microquake events is very small,leading to difficulty in locating microquake events.
Just as microquakes induced by hydraulic fracturing in a laboratory specimen can be monitored,large-scale microquakes induced by hydraulic fracturing in the field also can be monitored.And because the on-site monitoring region is large,the micro-quake source(hydraulic fracturing point)can be located at the same time.Therefore,microquake events induced by hydraulic fracturing can be monitored by a microseismograph during hydraulic fracturing in the field,le
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