CN102979579B - Method for analyzing coal and gas outburst risk in real time - Google Patents

Abstract

The invention relates to a method for analyzing coal and gas outburst risk in real time. The method comprises the steps of: first, arranging an air velocity sensor which is 20m far away from an inlet and outlet of an excavation roadway of a coal mine based on a current safety monitoring system of the coal mine; then, obtaining real time concentration of methane and real time air velocity of mixed air flow of a headwork face from monitoring data of the system; then, obtaining the sectional area of the headwork face and the cross sectional area of an air outlet of an hair dryer by consultation or measurement; then, calculating corresponding real time gas gushing amount per minute in the normal exploitation process according to the data; and finally, obtaining required data required according to the steps, and calculating coal and gas outburst risk analysis indexes with 15 days as a calculating period, wherein the outburst risk analysis indexes comprise crustal stress index id, gas pressure index ip, gas content index iw and coal desorption characteristic index ix.

Description

The method of real-time analysis coal and gas prominent danger

Technical field

The invention belongs to the technical field about driving face Coal and Gas Outbursts Prediction in safety of coal mines exploitation, a kind of particularly method of real-time analysis coal and gas prominent danger.

Background technology

Coal and gas prominent is the disaster affecting coal enterprises in China safety in production always, and along with the increase of mining depth, the harmfulness of this Accidents Disasters is increasing, has become the key factor of restriction coal in China industry development.Through the Researching and practicing of decades, Chinese scholars generally believes that coal and gas prominent is by the result of the physico-mechanical properties Three factors comprehensive function of geostatic stress, gas and coal, and define region and local " two quaternities " comprehensive outburst control measure technical system, require that first region protrusion-dispelling and local protrusion-dispelling will carry out outburst hazard prediction.Therefore, coal and gas prominent danger forecasting is the key of comprehensive outburst control measure technical system.

Existing outburst hazard Forecasting Methodology has three classes: 1. contact index prediction.Adopt drilling cuttings desorption of mash gas index K ₁or Δ h ₂, the index such as coal powder quantity of bore S and gas bearing capacity W predicts, the method is put into " control coal and gas prominent regulation ", becomes a kind of compulsive method of colliery scene extensive use.2. contactless concentration feature method.By analyzing the feature of outstanding front gas density curve, using parameters such as the Moving Average of concentration curve, amplitude, the frequency, variance and crest ratios as outburst hazard early warning foundation.3. contactless index prediction.The rules such as the outstanding Three factors variation characteristic of employing electromagnetic radiation, sound emission, gas density are analyzed, and achieve contactless prediction.Above-mentioned three class outburst prediction methods have played significant role in coal and gas outburst prominent controlling work, but also exist following main not enough at present: the drilling cuttings desorption of mash gas index K of 1. on-the-spot extensive use ₁or Δ h ₂, the contact index prediction method complex operation such as coal powder quantity of bore S, predict the outcome comparatively large by artificial affecting, the some effects factor of coal and gas prominent can only be reflected, and outburst hazard can not be analyzed continuously.2. utilize the parameters such as the Moving Average of gas density curve, variance to predict, lack theory support, prediction principle only rests on the empirical analysis stage.3. electromagnetic radiation and sound emission predicted method not very contactless prediction strictly speaking, can not realize real time continue analysis, and effect is undesirable in Coal and Gas Outbursts Prediction.4. utilize the outstanding Three factors variation characteristic of gas density to predict, prediction index has certain theoretical foundation, but does not consider that air quantity changes the impact brought predicting the outcome.In addition, its based on desorption of mash gas speed formula be applicable to loose coal, there is certain limitation, cause analysis result not accurate enough.5. the analysis indexes of often kind of Forecasting Methodology employing is no more than three, can not reflect the influence factor of multiple coal and gas prominent comprehensively, need increase analysis indexes, improves the accuracy of analysis result.

In addition, Safety Monitoring Control System of Coal Mine is one of coal mine downhole safety hedging " six Iarge-scale system ", the current whole nation substantially all collieries is all furnished with Safety Monitoring Control System of Coal Mine, achieve the real-time dynamic monitoring to coal mine gas, carbonomonoxide concentration, temperature, wind speed etc., but current monitoring and controlling system only achieves monitoring and the display of all kinds of parameter, still lack the abundant analysis to monitored data and utilization.

Summary of the invention

The technical problem to be solved in the present invention be propose a kind ofly to utilize existing Safety Monitoring Control System of Coal Mine, real time continue analysis can be realized, effect better, the method for analytical structure real-time analysis coal and gas prominent danger more accurately.

The technical scheme realizing the object of the invention is to provide a kind of method of real-time analysis coal and gas prominent danger, comprises the steps:

1. Safety Monitoring Control System of Coal Mine is built, arrange at the position being less than or equal to 5m at the distance driving face place of the digging laneway of coal road and be arranged in the first distinguished and admirable methane transducer of mixing, arrange at 10 to the 15m place of the import and export of distance digging laneway and be arranged in the second distinguished and admirable methane transducer of mixing, booster is set being arranged in the tunnel adjacent with digging laneway, and the air outlet of the external air duct of this booster is positioned at the position near driving face, the air intake of this booster is then arranged on booster, and booster air intake towards 3 to the 5m place in direction, leucoaurin sensor is set.In digging laneway and 18 to the 22m place of import and export of distance digging laneway arranges an air velocity transducer.

2. from the monitored data of Safety Monitoring Control System of Coal Mine, obtain the Real-time Monitoring Data of the first methane transducer in digging laneway and air velocity transducer, and using the data obtained as the distinguished and admirable real-time methane concentration data C of the mixing of driving face and wind speed size data V in real time, namely, when the time is t, corresponding methane concentration data are C _t, real-time wind speed size data V _t.

3. by consulting the design data of the driving face in colliery or being obtained the cross-sectional area S of the driving face that digging laneway designs by the method measured ₁with the cross sectional area S of the air outlet of air duct ₂.

4. according to the data that 2. and 3. step obtains, the real-time gas emission Q of following formulae discovery driving face in normal recovery process is utilized _t:

Q _t=C _tv _t(S ₁-S ₂), wherein t is the corresponding time.

5. 2., 3. and 4. obtain required data according to step, calculate every coal and gas prominent risk analysis index, outburst hazard analysis indexes comprises geostatic stress index i _d, gas pressure index i _p, gas bearing capacity index i _wwith coal body desorption properties index i _x, the computing cycle of above-mentioned each outburst hazard analysis indexes is 15 days, represents the sky number sequence value in computing cycle, r=1,2, L, 15 with r:

A geostatic stress index i _d:

Select 15 days computing cycles as day moving average, according to the gas emission size that 4. step obtains, utilize the day outburst amount average Q of the every day in this computing cycle of 15 days of following formulae discovery _prand the day moving average Q in this computing cycle of 15 days _y:

$Q_{\mathrm{pr}}=\frac{1}{24}\underset{t=1}{\overset{24}{\mathrm{\Σ}}}{Q}_{\mathrm{tr}}.$

$Q_y=\frac{1}{360}\underset{t=1}{\overset{360}{\mathrm{\Σ}}}{Q}_t.$

In above-mentioned formula, Q _trrefer to the hourly real-time gas emission in r days 24 hours, Q _trefer to the gas emission hourly in this computing cycle of 15 days.Then formula is utilized the geostatic stress index of every day that can calculate this driving face in the computing cycle of above-mentioned 15 days.

B gas pressure index i _p:

In the normal recovery process in colliery, the size of coal gas outburst amount constantly fluctuates along with passage of time, reaches peak value when driving face coal breakage starts.Gas cumulative amount Q' in the normal recovery process in colliery can adopt formula Q'=Alnt'+B to describe, and wherein t' is the time, and A, B refer to index of correlation, and lnt' refers to that t' takes the logarithm.Using the coefficient A in formula Q'=Alnt'+B as gas pressure index i _p.

The gas emission peak value of the maximum gas emission peak value of numerical value within 24 hours same day as the same day is chosen in every day, using the time t corresponding to the gas emission peak value that is selected as the gas emission peak value on the same day as start time t'=0, calculated to the Gas cumulative amount of corresponding time point a period of time in 20 to 30min after time point corresponding to the gas emission peak value be selected as the gas emission peak value on the same day at interval of 1 to 3 minute, take t' as abscissa, Gas cumulative amount Q' is ordinate, make Gas cumulative amount curve, be abscissa according to Gas cumulative amount curve with lnt' again, Gas cumulative amount Q' is the relation curve that ordinate makes Gas cumulant Q' and the lnt' in the time period of this 20 to 30min, the relation curve of the linear fit instrument in Excel to above-mentioned Gas cumulant Q' and lnt' is utilized to carry out matching, the numerical value of the coefficient A in formula Q'=Alnt'+B is obtained after matching, as the index i of gas pressure characterizing the same day _p, calculate the gas pressure index i of every day _p.

C gas bearing capacity index i _w:

The day rolling average outburst amount Q in the computing cycle of 15 days has been calculated in a _y, choose the gas emission peak value Q of a maximum gas emission peak value of numerical value within 24 hours same day as the same day in every day _fr, utilize formula the gas bearing capacity index of every day that can calculate driving face within this time period of 15 days.

D coal body desorption properties index i _x:

Gas cumulative amount Q' in the normal recovery process in colliery also can be described by following 2 relational expressions:

$Q^{\′}=V_0^{\′}\left[\frac{{(1+t^{\′})}^{1-n}-1}{1-n}\right],$

Q'＝a·t' ⁱ，

T' refers to the time.V ₀' Gas speed when referring to t'=0.A, i, n refer to the constant relevant with the gas bearing capacity of coal and structure.

The gas emission peak value of the maximum gas emission peak value of numerical value within 24 hours same day as the same day is chosen in every day, using the time t corresponding to the gas emission peak value that is selected as the gas emission peak value on the same day as start time t'=0, calculated to the Gas cumulative amount of corresponding time point a period of time in 20 to 30min after time point corresponding to the gas emission peak value be selected as the gas emission peak value on the same day at interval of 1 to 3 minute, take t' as abscissa, Gas cumulative amount Q' is that ordinate makes Gas cumulative amount curve, the lsqcurvefit function in Matlab is utilized to carry out matching to above-mentioned Gas cumulative amount curve, formula can be obtained in the numerical value of coefficient n and Q'=at' ⁱthe numerical value of Exponential i, using n and i all as the index i characterizing coal body desorption properties _x.

Step 5. in, calculate b gas pressure index i _ptime, the derivation of Gas cumulative amount Q'=Alnt'+B is:

Coal gas desorb cumulative amount and time take the logarithm after value there is following linear relationship: Q _j=Alnt'+B, in above-mentioned formula, Q _jrefer to desorption of mash gas cumulative amount.T' refers to the time.A, B refer to index of correlation.Gas pressure is reflected in the difference of coefficient A for the impact of coal gas desorption properties in logarithmic formula, and gas pressure is larger, then faster with desorption rate in the time period, desorption of mash gas amount growth trend is more obvious, and A is larger for its fitting coefficient.

In the normal recovery process in colliery, the size of coal gas outburst amount constantly fluctuates along with passage of time, reaches peak value when driving face coal breakage starts.And in the normal recovery process in colliery, the coal breakage amount of driving face is substantially constant, in the moment that coal breakage starts, the gas emission of driving face reaches maximum, and afterwards along with dilution distinguished and admirable in coal breakage process, gas emission diminishes gradually.Process due to driving face coal breakage is equivalent to the process of coal body desorb, therefore in the Gas cumulative amount of coal breakage process and coal body desorption process, desorption of mash gas cumulative amount can relative indicatrix each other each other, therefore, coal gas desorb cumulative amount and the relation between the time are applicable to Gas cumulative amount and the relation between the time.Because coal gas desorb cumulative amount and the relation between the time are also applicable to Gas cumulative amount and the relation between the time, then Gas cumulative amount and time take the logarithm after value also there is following linear relationship: Gas cumulative amount Q'=Alnt'+B.

Step 5. in, calculate d coal body desorption properties index i _xtime, Gas cumulative amount $Q^{\′}={V^{\′}}_0\left[\frac{{(1+t^{\′})}^{1-n}-1}{1-n}\right]$ And Q'=at' ⁱderivation be:

Based on ickings gas Diffusion Law, the Wu Si Jino husband's formula in existing desorption of mash gas formula is:

$Q_j=V_0\left[\frac{{(1+t^{\′})}^{1-n}-1}{1-n}\right],$

Sun Chong rising sun formula is:

Q _j＝a·t' ⁱ，

In above-mentioned formula, Q _jrefer to desorption of mash gas cumulative amount.T' refers to the time.V ₀refer to desorption of mash gas speed during t'=0.A, i, n refer to the constant relevant with the gas bearing capacity of coal and structure.

In the normal recovery process in colliery, the size of coal gas outburst amount constantly fluctuates along with passage of time, reaches peak value when driving face coal breakage starts.And in the normal recovery process in colliery, the coal breakage amount of driving face is substantially constant, in the moment that coal breakage starts, the gas emission of driving face reaches maximum, and afterwards along with dilution distinguished and admirable in coal breakage process, gas emission diminishes gradually.Process due to driving face coal breakage is equivalent to the process of coal body desorb, therefore in the Gas cumulative amount of coal breakage process and coal body desorption process, desorption of mash gas cumulative amount can relative indicatrix each other each other, then Gas cumulative amount Q' also can be described by following relational expression:

$Q^{\′}=V_0^{\′}\left[\frac{{(1+t^{\′})}^{1-n}-1}{1-n}\right],$ V ₀' Gas speed when referring to t'=0,

Q'＝a·t' ⁱ。

Step 1. in, before and after air velocity transducer, 10m planted agent is distinguished and admirable without branch, nothing is turned round, accessible, section is unchanged.

Leucoaurin sensor is for detecting in the air-flow of its position the data of contained methane, and these data are as the data of methane contained by driving face air intake.

The technical scheme realizing the object of the invention is to provide the present invention and has positive effect: the monitored data that the method for (1) real-time analysis of the present invention coal and gas prominent danger utilizes existing coal mine safety monitoring system to monitor can the outburst hazard of real time continue analysis driving face, time saving and energy saving, the precision of prediction of various parameter is higher.

(2) Forecasting Methodology of original contact is improved to present contactless analytical method by the method for real-time analysis coal and gas prominent danger of the present invention, achieve the outburst hazard of online in-situ analysis driving face, become successional analysis from the test of discontinuous.

(3) original hand drilling prediction is improved to present system automatic analysis by the method for real-time analysis coal and gas prominent danger of the present invention, eliminates the impact of the human factors such as manual operation, does not affect production, make test result more accurate.

(4) method of real-time analysis coal and gas prominent danger of the present invention considers air quantity and changes the impact brought predicting the outcome, increase and optimize dynamic prediction index, the customer service defect of original real-time predicting method, makes analysis result more closer to reality.

(5) method of real-time analysis coal and gas prominent danger of the present invention only needs increase to arrange an air velocity transducer, and air velocity transducer is much lower relative to desorption instrument, the cost such as to drill, and significantly reduces forecast analysis expense.

Accompanying drawing explanation

Fig. 1 be the Safety Monitoring Control System of Coal Mine of the driving face of series ventilation in colliery schematic diagram is set;

Fig. 2 be in embodiment 1 certain year certain month 16 days from 0: 0 assign to 23: 59 on the 30th during this period of time in the normal recovery process in colliery the gas density of driving face and wind speed curve;

Fig. 3 be colliery in embodiment 1 driving face certain year certain month 16 days from 0: 0 assign to 23: 59 on the 30th during this period of time in Gas curve;

Fig. 4 be colliery in embodiment 1 driving face certain year certain month 16 days from 0: 0 assign to 23: 59 on the 30th during this period of time in geostatic stress index, gas pressure index, gas bearing capacity index and coal body desorption properties index change curve separately and the traditional drilling cuttings desorption of mash gas index K for contrasting ₁change curve;

Fig. 5 be colliery in embodiment 1 driving face described certain year certain month 16 days from 0: 0 assign to 23: 59 on the 30th during this period of time in the Gas curve of 22 days;

Gas emission in the 26min between 54min to 79min that Fig. 6 is the Gas curve to Fig. 5 adds up obtained gas emission accumulation curve;

Fig. 7 is the relation curve of Gas cumulant and the lnt' made according to the gas emission accumulation curve of Fig. 6.

Mark in above-mentioned accompanying drawing is as follows:

Digging laneway B, driving face B1, import and export B2, tunnel A,

First methane transducer T ₁, the second methane transducer T ₂, leucoaurin sensor T ₃, air velocity transducer S, booster F, air intake F1, air outlet F2,

The cross-sectional area S of digging laneway design ₁, air duct cross sectional area S ₂.

Detailed description of the invention

(embodiment 1)

The present embodiment is for the process of a certain driving face enforcement real-time analysis coal and gas prominent danger in Shanxi Lu An group colliery, and the method for the real-time analysis coal and gas prominent danger of the present embodiment comprises following several step:

1. Safety Monitoring Control System of Coal Mine is built.See Fig. 1, the driving face B1 of the present embodiment belongs to the work plane of the digging laneway B of series ventilation coal road, according to " coal mine safety monitoring system and detecting instrument use management specification " 6.4 joints 6.4.1 in coal road, half coal petrography and have the requirement of design code of driving face methane transducer of Gas rock gangway, arrange to be arranged at the position being less than or equal to 5m at the distance driving face B1 place of the digging laneway B of coal road and mix the first distinguished and admirable methane transducer T ₁, arrange at 10 to the 15m place at the import and export B2 place of distance digging laneway B and be arranged in the second distinguished and admirable methane transducer T of mixing ₂owing to adopting the driving face of series ventilation, also arrange and gone here and there work plane booster F being arranged in the tunnel A adjacent with digging laneway B, and the air outlet F2 of the external air duct of this booster F is positioned at the position near driving face B1, the air intake F1 of this booster F is then arranged on booster F, and booster F air intake F1 towards 3 to the 5m place in direction, leucoaurin sensor T is set ₃, leucoaurin sensor T ₃for detecting in the air-flow of its position the data of contained methane, these data are as the data of methane contained by driving face air intake.Methane transducer T ₁, T ₂, T ₃answer vertical hanging, must not 0.3m be greater than apart from top board, must not 0.2m be less than apart from lane wall.The methane transducer mixing distinguished and admirable place should have the protective equipment preventing blasting impact.

Still see Fig. 1, in the present embodiment, Safety Monitoring Control System of Coal Mine also arranges an air velocity transducer S in the 20m place increase of the import and export B2 of the digging laneway B in distance colliery, and before and after air velocity transducer S, 10m planted agent is distinguished and admirable without branch, nothing is turned round, accessible, section is unchanged.

2. from the monitored data of Safety Monitoring Control System of Coal Mine, obtain the first methane transducer T in digging laneway B ₁with the Real-time Monitoring Data of air velocity transducer S, and using the data obtained as the distinguished and admirable real-time methane concentration data C of the mixing of driving face B1 and wind speed size data V in real time, when namely the time is t, corresponding methane concentration data are C _t, real-time wind speed size data V _t.

In the present embodiment, driving face B1 to be assigned in 23: 59 on the 30th these 15 days from 0: 0 at certain year certain month 16 days and gathers the methane concentration C hourly of this development end B1 and the Real-time Monitoring Data of wind speed size V, as shown in Figure 2, using 0: 0 on the 16th as time zero, gas density and the wind speed curve of this driving face B1 is drawn according to above-mentioned Real-time Monitoring Data, as seen from the figure, when 21h, methane concentration C _21hsize be 0.336% (gas percent concentration), real-time wind speed V _21hsize be 1.06m/s; When 220h, methane concentration C _220hsize be 0.434%, real-time wind speed V _220hsize be 1.11m/s.

3. Fig. 1 is seen, by consulting the design data of the driving face B1 in colliery or being obtained the cross-sectional area S of the driving face B1 that digging laneway designs by the method measured ₁with the cross sectional area S of the air outlet F2 of air duct ₂, unit is m ².In the present embodiment, the cross-sectional area S of the design of the driving face B1 of the digging laneway B in this colliery ₁for 13.5m ², the diameter of the air outlet F2 of air duct is 500mm, the cross sectional area S of the air outlet F2 of air duct ₂for 0.2m ².

4. according to the data that 2. and 3. step obtains, the real-time gas emission Q of following formulae discovery driving face B1 within the above-mentioned time period is utilized _t, unit is m ³/ min:

Q _t=C _tv _t(S ₁-S ₂), wherein t is the corresponding time.

Such as, when time 21h, the C 2. obtained by step _21hsize be 0.336%, V _21hsize be 1.06m/s, thus by formulae discovery 21h time gas emission Q _21h, unit is m ³/ min;

Q _21h＝C _21h·V _21h(S ₁-S ₂)＝0.336％×1.06×(13.5-0.2)＝0.047m ³/s＝2.84m ³/min；

And for example when time 220h, the C 2. obtained by step _220hsize be 0.434%, V _220hsize be 1.11m/s, thus by formulae discovery 220h time gas emission Q ₂₂₀:

Q _220h＝C _220h·V _220h(S ₁-S ₂)＝0.434％×1.11×(13.5-0.2)＝0.064m ³/s＝3.84m ³/min；

The real-time gas emission of driving face B1 within the above-mentioned corresponding time period can be calculated according to the method described above, then according to the real-time gas emission Q calculated _tsize and corresponding time value draw obtain the Gas curve of driving face B1 within this time period of 15 days, as shown in Figure 3.

5. 2., 3. and 4. obtain required data according to step, calculate every coal and gas prominent risk analysis index, outburst hazard analysis indexes comprises geostatic stress index i _dgas pressure index i _p, gas bearing capacity index i _wwith coal body desorption properties index i _x, the computing cycle of above-mentioned each outburst hazard analysis indexes is 15 days, represents the sky number sequence value in computing cycle with r, in the present embodiment, and r=1,2, L, 15, such as, during r=3, be the 3rd day:

A geostatic stress index i _d:

Select 15 days computing cycles as day moving average, the gas emission size according to Fig. 3, utilize the day outburst amount average Q of the every day in this time period of following formulae discovery _prand the day moving average Q of this time period _y:

$Q_{\mathrm{pr}}=\frac{1}{24}\underset{t=1}{\overset{24}{\mathrm{\Σ}}}{Q}_{\mathrm{tr}};$

$Q_y=\frac{1}{360}\underset{t=1}{\overset{360}{\mathrm{\Σ}}}{Q}_t=3.49m^3/\min ;$

In above-mentioned formula, Q _trrefer to the hourly real-time gas emission in r days 24 hours, unit is m ³/ min; Q _trefer to the hourly real-time gas emission in this computing cycle of 15 days, unit is m ³/ min.Such as during r=3, calculate the day outburst amount average Q of the 3rd day (the 48 to 72 hour) _p3:

$Q_{p3}=\frac{1}{24}\underset{t=1}{\overset{24}{\mathrm{\Σ}}}{Q}_{t3}=3.642m^3/\min ;$

During r=12, calculate the day outburst amount average Q of the 12nd day (the 276 to 288 hour) _p12:

$Q_{p12}=\frac{1}{24}\underset{t=1}{\overset{24}{\mathrm{\Σ}}}{Q}_{t12}=3.867m^3/\min ;$

In the time period that can calculate corresponding 15 days according to the method described above, the average of gushing out day of every day, utilizes formula the geostatic stress index of every day that can calculate this driving face B1 within the above-mentioned time period.

Such as calculate the geostatic stress index i of the 3rd day _d3with the geostatic stress index i of the 12nd day _d12:

$i_{d3}=\frac{Q_{p3}}{Q_y}=\frac{3.642}{3.49}=1.04;$

$i_{d12}=\frac{Q_{p12}}{Q_y}=\frac{3.867}{3.49}=1.11;$

The geostatic stress index curve in the described time period is made in, as shown in Figure 4 after the geostatic stress index of every day that draws this driving face B1 within the above-mentioned time period.

B gas pressure index i _p:

Now there are some researches show, coal gas desorb cumulative amount and time take the logarithm after value there is good linear relationship (see page content of the 84th in document 1, [1] Fenghua is pacified, Cheng Yuanping, Wu Dongmei etc., " calculating the method [J] of coal-bed gas pressure based on desorption of mash gas characteristic " mining and safety first engineering journal, 2011,28 (1): 81-85,89), that is: Q _j=Alnt'+B, in above-mentioned formula, Q _jrefer to desorption of mash gas cumulative amount, unit is m ³/ min; T' refers to the time, and unit is min; A, B refer to index of correlation.Gas pressure is reflected in the difference of coefficient A for the impact of coal gas desorption properties in logarithmic formula, gas pressure is larger, then faster with desorption rate in the time period, desorption of mash gas amount growth trend is more obvious, its fitting coefficient A is larger (see page content of the 84th in document 1, [1] Fenghua is pacified, Cheng Yuanping, Wu Dongmei etc., " calculating the method [J] of coal-bed gas pressure based on desorption of mash gas characteristic " mining and safety first engineering journal, 2011,28 (1): 81-85,89).

In the normal recovery process in colliery, the size of coal gas outburst amount constantly fluctuates along with passage of time, reaches peak value when driving face B1 coal breakage starts.And in the normal recovery process in colliery, the coal breakage amount of driving face B1 is substantially constant, in the moment that coal breakage starts, the gas emission of driving face B1 reaches maximum, and afterwards along with dilution distinguished and admirable in coal breakage process, gas emission diminishes gradually.Process due to driving face B1 coal breakage is equivalent to the process of coal body desorb, therefore in the Gas cumulative amount of coal breakage process and coal body desorption process, desorption of mash gas cumulative amount can relative indicatrix each other each other, therefore, coal gas desorb cumulative amount and the relation between the time are applicable to Gas cumulative amount and the relation between the time.Because coal gas desorb cumulative amount and the relation between the time are also applicable to Gas cumulative amount and the relation between the time, then Gas cumulative amount and time take the logarithm after value also there is good linear relationship, therefore, Gas cumulative amount Q' also can be described by above-mentioned relational expression: Q'=Alnt'+B, using the coefficient A in formula as gas pressure index i _p.Choose a gas emission peak value below to calculate, and be illustrated.

Such as, this gas emission data analysis of 22 days in 15 days calculates to choose certain year certain month 16 to 30 of the present embodiment, the Gas discharge curve of 22 days as shown in Figure 5, as seen from Figure 5, the size of gas emission along with passage of time be constantly fluctuation, occur multiple peak value, the beginning of the corresponding coal breakage of each peak value, in order to calculate gas pressure index i _p, need to make the Gas cumulative amount curve that coal breakage starts a period of time in rear 20 to 30min, choose the gas emission peak value of the maximum gas emission peak value of numerical value within 24 hours same day on the 22nd as the same day, as seen from Figure 5, within 22 days 0: 54, be divided into the maximum gas emission peak value of numerical value on the same day, Gas cumulative amount curve is made at interval of the Gas cumulative amount calculating corresponding time point for 2 minutes from 0: 54 on the 22nd, Fig. 6 is the Gas cumulative amount curve in the 26min after starting for 0: 54 on the 22nd, the relation curve of Gas cumulant and lnt' in this time period is made again according to Gas cumulative amount curve, as shown in Figure 7, the linear fit instrument in Excel is utilized to carry out matching to the curve in Fig. 7, its matching degree of agreement R ²=0.9688, close to 1, the matching goodness of fit is higher, obtains the coefficient A=20.87 in formula Q'=Alnt'+B after matching, as the index i of gas pressure characterizing the same day _p, i.e. the gas pressure index i of 22 days _p=A=20.87.

Therefore, the gas emission peak value of the maximum gas emission peak value of numerical value within 24 hours same day as the same day can be chosen according to the method described above in every day, using the time t corresponding to the gas emission peak value that is selected as the gas emission peak value on the same day as start time t'=0, calculated to the Gas cumulative amount of corresponding time point a period of time in 20 to 30min after time point corresponding to the gas emission peak value be selected as the gas emission peak value on the same day at interval of 2 minutes, take t' as abscissa, Gas cumulative amount Q' is ordinate, make Gas cumulative amount curve, be abscissa according to Gas cumulative amount curve with lnt' again, Gas cumulative amount Q' is the relation curve that ordinate makes Gas cumulant Q' and lnt' in this time period, the relation curve of the linear fit instrument in Excel to above-mentioned Gas cumulant Q' and lnt' is utilized to carry out matching, the numerical value of the coefficient A in formula Q'=Alnt'+B is obtained after matching, as the index i of gas pressure characterizing the same day _p, calculate the gas pressure index i of the every day in the time period of 15 days of the present embodiment _p, the gas pressure index of 15 days according to the present embodiment makes gas pressure index curve, and result as shown in Figure 4.

C gas bearing capacity index i _w:

The day rolling average outburst amount Q in this time period of 15 days has been calculated in a _yfor 3.49m ³/ min, chooses the gas emission peak value Q of a maximum gas emission peak value of numerical value within 24 hours same day as the same day in every day _fr, the gas emission peak value Q of such as the 3rd day _f3for 4.222m ³/ min, the gas emission peak value Q of the 12nd day _f12for 4.41m ³/ min.Utilize empirical formula the gas bearing capacity index of every day that can calculate driving face B1 within this time period of 15 days, such as:

$i_{w3}=\frac{Q_{f3}}{Q_y}=\frac{4.222}{3.49}=1.21,$

$i_{w12}=\frac{Q_{f12}}{Q_y}=\frac{4.41}{3.49}=1.26.$

The gas bearing capacity index curve in the described time period is made in, as shown in Figure 4 after the gas bearing capacity index of every day that draws this driving face B1 within the time period of 15 days of the present embodiment.

D coal body desorption properties index i _x:

The Coal and Gas Outbursts Prediction index great majority that in " control coal and gas prominent specifies ", regulation adopts are relevant to coal body desorption properties, and coal body desorption properties can reflect this factor of physico-mechanical properties of coal well.Based on ickings gas Diffusion Law, Chinese scholars proposes multiple desorption of mash gas formula, and wherein Wu Si Jino husband formula and the application of Sun Chong rising sun formula are comparatively extensively, as follows respectively:

$Q_j=V_0\left[\frac{{(1+t^{\′})}^{1-n}-1}{1-n}\right],$

Q _j＝a·t' ⁱ，

In above-mentioned formula, Q _jrefer to desorption of mash gas cumulative amount, unit is m ³/ min; T' refers to the time, and unit is min; V ₀refer to desorption of mash gas speed during t'=0, unit is cm ³/ (gmin); A, i, n refer to the constant relevant with the gas bearing capacity of coal and structure.

Because desorption of mash gas cumulative amount in the Gas cumulative amount of coal breakage process and coal body desorption process can relative indicatrix each other each other, therefore, Gas cumulative amount Q' also can be described by following relational expression:

$Q^{\′}=V_0^{\′}\left[\frac{{(1+t^{\′})}^{1-n}-1}{1-n}\right],$ V ₀' Gas speed when referring to t'=0,

Q'＝a·t' ⁱ，

But these two formula are also not suitable for the desorption properties of all ickings, even can there is larger error in some ickings adsorption laws of description.Therefore, can matching be carried out with two formula simultaneously thus obtain the formula of applicable coal sample adsorption law, using coefficient n or i as sign coal body desorption properties index i _x.Choose a gas emission peak value below to calculate, and be illustrated.

Such as, in Figure 5, the size of gas emission is constantly fluctuation along with passage of time, occurs multiple gas emission peak value, and the beginning of the corresponding coal breakage of each gas emission peak value, in order to calculate coal body desorption properties index i _x, need to make the Gas cumulative amount curve that coal breakage starts a period of time in rear 20 to 30min, choose the gas emission peak value of the maximum gas emission peak value of numerical value within 24 hours same day on the 22nd as the same day, as seen from Figure 5, within 22 days 0: 54, be divided into the maximum gas emission peak value of numerical value on the same day, Gas cumulative amount curve is made at interval of the Gas cumulative amount calculating corresponding time point for 2 minutes from 0: 54 on the 22nd, Fig. 6 is the Gas cumulative amount curve in the 26min after starting for 0: 54 on the 22nd, the lsqcurvefit function in Matlab is utilized to carry out matching to above-mentioned Gas cumulative amount curve, formula can be obtained in the size of coefficient n be 0.25 and Q'=at' ⁱthe size of Exponential i is 0.69, as the New Set of the coal body desorption properties on sign same day on the 22nd.

The gas emission peak value of the maximum gas emission peak value of numerical value within 24 hours same day as the same day can be chosen according to the method described above in every day, using the time t corresponding to the gas emission peak value that is selected as the gas emission peak value on the same day as start time t'=0, calculated to the Gas cumulative amount of corresponding time point a period of time in 20 to 30min after time point corresponding to the gas emission peak value be selected as the gas emission peak value on the same day at interval of 2 minutes, take t' as abscissa, Gas cumulative amount Q' is ordinate, make Gas cumulative amount curve, the lsqcurvefit function in Matlab is utilized to carry out matching to above-mentioned Gas cumulative amount curve, formula can be obtained in the numerical value of coefficient n and Q'=at' ⁱthe numerical value of Exponential i, as the index i of coal body desorption properties characterizing the same day _x.The coal body desorption properties index n of every day and i in the time period of 15 days calculating the present embodiment, and make the curve of gas bearing capacity index n and i respectively, as shown in Figure 4, n and i all can as coal body desorption properties index i _x.

To the desorption of mash gas index K with the actual measurement of traditional colliery scene in recovery process within the time period of above-mentioned 15 days of the driving face B1 of the present embodiment ₁assay method measures, and makes K ₁change curve, the outburst hazard analysis indexes 5. calculated with step contrasts, and as shown in Figure 4, can find out geostatic stress index i _d, gas pressure index i _p, gas bearing capacity index i _wwith variation tendency and the K of the curve of these four indexs of coal body desorption properties index n ₁the variation tendency of curve is identical, such as, when r=6 (the 6th day), and K ₁=0.385, i _d=1.183, i _p=22.1, i _w=1.472, n=0.34, is maximum; When r=12 (the 12nd day), K ₁=0.18, i _d=0.902, i _p=18.23, i _w=0.972, n=0.18, is minimum.The variation tendency of the curve of coal body desorption properties index i and K ₁the variation tendency of curve is just in time contrary, such as, when r=6, and K ₁=0.385, be maximum, and i=0.531, be minimum; When r=12, K ₁=0.18, be minimum, and i=0.77, be maximum.The above results shows the outburst hazard analysis indexes that the present invention calculates and traditional desorption of mash gas index K ₁there is synchronism, by contrast K ₁outburst danger threshold, the threshold of these outburst danger analysis indexes can be obtained, for coal and gas prominent work plane prediction foundation is provided.

Claims (4)

1. a method for real-time analysis coal and gas prominent danger, is characterized in that, comprises the steps:

1. build Safety Monitoring Control System of Coal Mine, arrange at the position being less than or equal to 5m at distance driving face (B1) place of the digging laneway (B) of coal road and be arranged in the first distinguished and admirable methane transducer (T of mixing ₁), 10 to the 15m place at import and export (B2) place of distance digging laneway (B) arranges and is arranged in the second distinguished and admirable methane transducer (T of mixing ₂), booster (F) is set being arranged in the tunnel (A) adjacent with digging laneway (B), and the air outlet (F2) of the external air duct of this booster (F) is positioned at the position near driving face (B1), the air intake (F1) of this booster (F) is then arranged on booster (F), and booster (F) air intake (F1) towards 3 to the 5m place in direction, leucoaurin sensor (T is set ₃); In digging laneway (B) and 18 to the 22m place of import and export (B2) of distance digging laneway (B) arranges an air velocity transducer (S);

2. from the monitored data of Safety Monitoring Control System of Coal Mine, obtain the Real-time Monitoring Data of the first methane transducer (T1) in digging laneway (B) and air velocity transducer (S), and using the data obtained as the distinguished and admirable real-time methane concentration data C of the mixing of driving face (B1) and wind speed size data V in real time, namely, when the time is t, corresponding methane concentration data are C _t, real-time wind speed size data V _t;

3. by consulting the design data of the driving face (B1) in colliery or being obtained the cross-sectional area S of the driving face (B1) that digging laneway designs by the method measured ₁with the cross sectional area S of the air outlet (F2) of air duct ₂;

4. according to the data that 2. and 3. step obtains, the real-time gas emission Q of following formulae discovery driving face (B1) in normal recovery process is utilized _t:

Q _t=C _tv _t(S ₁-S ₂), wherein t is the corresponding time;

5. 2., 3. and 4. obtain required data according to step, calculate every coal and gas prominent risk analysis index, outburst hazard analysis indexes comprises geostatic stress index i _d, gas pressure index i _p, gas bearing capacity index i _wwith coal body desorption properties index i _x, the computing cycle of above-mentioned each outburst hazard analysis indexes is 15 days, represents the sky number sequence value in computing cycle, r=1,2, L, 15 with r:

A geostatic stress index i _d:

Select 15 days computing cycles as day moving average, according to the gas emission size that 4. step obtains, utilize the day outburst amount average Q of the every day in this computing cycle of 15 days of following formulae discovery _prand the day moving average Q in this computing cycle of 15 days _y:

$Q_{\mathrm{pr}}=\frac{1}{24}\underset{t=1}{\overset{24}{\mathrm{\Σ}}}{Q}_{\mathrm{tr}}$

$Q_y=\frac{1}{360}\underset{t=1}{\overset{360}{\mathrm{\Σ}}}{Q}_t;$

In above-mentioned formula, Q _trrefer to the hourly real-time gas emission in r days 24 hours, Q _trefer to the gas emission hourly in this computing cycle of 15 days; Then formula is utilized the geostatic stress index of every day that can calculate this driving face (B1) in the computing cycle of above-mentioned 15 days;

B gas pressure index i _p:

In the normal recovery process in colliery, the size of coal gas outburst amount constantly fluctuates along with passage of time, reaches peak value when driving face (B1) coal breakage starts; Gas cumulative amount Q' in the normal recovery process in colliery adopts formula Q'=Alnt'+B to describe, and wherein t' is the time, and A, B refer to index of correlation, and lnt' refers to that t' takes the logarithm; Using the coefficient A in formula Q'=Alnt'+B as gas pressure index i _p;

The gas emission peak value of the maximum gas emission peak value of numerical value within 24 hours same day as the same day is chosen in every day, using the time t corresponding to the gas emission peak value that is selected as the gas emission peak value on the same day as start time t'=0, calculated to the Gas cumulative amount of corresponding time point a period of time in 20 to 30min after time point corresponding to the gas emission peak value be selected as the gas emission peak value on the same day at interval of 1 to 3 minute, take t' as abscissa, Gas cumulative amount Q' is ordinate, make Gas cumulative amount curve, be abscissa according to Gas cumulative amount curve with lnt' again, Gas cumulative amount Q' is the relation curve that ordinate makes Gas cumulant Q' and the lnt' in the time period of this 20 to 30min, the relation curve of the linear fit instrument in Excel to above-mentioned Gas cumulant Q' and lnt' is utilized to carry out matching, the numerical value of the coefficient A in formula Q'=Alnt'+B is obtained after matching, as the index i of gas pressure characterizing the same day _p, calculate the gas pressure index i of every day _p,

C gas bearing capacity index i _w:

The day rolling average outburst amount Q in the computing cycle of 15 days has been calculated in a _y, choose the gas emission peak value Q of a maximum gas emission peak value of numerical value within 24 hours same day as the same day in every day _fr, utilize formula the gas bearing capacity index of every day that can calculate driving face (B1) within this time period of 15 days;

D coal body desorption properties index i _x:

Following 2 relational expressions of Gas cumulative amount Q' in the normal recovery process in colliery are described:

$Q^{\′}=V_0^{\′}\left[\frac{{(1+t^{\′})}^{1-n}-1}{1-n}\right],$

Q'＝a·t' ⁱ，

T' refers to the time; V ₀' Gas speed when referring to t'=0; A, i, n refer to the constant relevant with the gas bearing capacity of coal and structure;

The gas emission peak value of the maximum gas emission peak value of numerical value within 24 hours same day as the same day is chosen in every day, using the time t corresponding to the gas emission peak value that is selected as the gas emission peak value on the same day as start time t'=0, calculated to the Gas cumulative amount of corresponding time point a period of time in 20 to 30min after time point corresponding to the gas emission peak value be selected as the gas emission peak value on the same day at interval of 1 to 3 minute, take t' as abscissa, Gas cumulative amount Q' is that ordinate makes Gas cumulative amount curve, the lsqcurvefit function in Matlab is utilized to carry out matching to above-mentioned Gas cumulative amount curve, formula can be obtained in the numerical value of coefficient n and Q'=at' ⁱthe numerical value of Exponential i, using n and i all as the index i characterizing coal body desorption properties _x.

2. the method for real-time analysis coal and gas prominent danger according to claim 1, is characterized in that: step 5. in, calculate b gas pressure index i _ptime, the derivation of Gas cumulative amount Q'=Alnt'+B is:

Coal gas desorb cumulative amount and time take the logarithm after value there is following linear relationship: Q _j=Alnt'+B, in above-mentioned formula, Q _jrefer to desorption of mash gas cumulative amount; T' refers to the time; A, B refer to index of correlation; Gas pressure is reflected in the difference of coefficient A for the impact of coal gas desorption properties in logarithmic formula, and gas pressure is larger, then faster with desorption rate in the time period, desorption of mash gas amount growth trend is more obvious, and A is larger for its fitting coefficient;

In the normal recovery process in colliery, the size of coal gas outburst amount constantly fluctuates along with passage of time, reaches peak value when driving face (B1) coal breakage starts; And in the normal recovery process in colliery, the coal breakage amount of driving face (B1) is substantially constant, and in the moment that coal breakage starts, the gas emission of driving face (B1) reaches maximum, afterwards along with dilution distinguished and admirable in coal breakage process, gas emission diminishes gradually; Process due to driving face (B1) coal breakage is equivalent to the process of coal body desorb, therefore in the Gas cumulative amount of coal breakage process and coal body desorption process, desorption of mash gas cumulative amount can relative indicatrix each other each other, therefore, coal gas desorb cumulative amount and the relation between the time are applicable to Gas cumulative amount and the relation between the time; Because coal gas desorb cumulative amount and the relation between the time are also applicable to Gas cumulative amount and the relation between the time, then Gas cumulative amount and time take the logarithm after value also there is following linear relationship: Gas cumulative amount Q'=Alnt'+B.

3. the method for real-time analysis coal and gas prominent danger according to claim 1, is characterized in that: step 5. in, calculate d coal body desorption properties index i _xtime, Gas cumulative amount and Q'=at' ⁱderivation be:

Based on ickings gas Diffusion Law, the Wu Si Jino husband's formula in existing desorption of mash gas formula is:

$Q_j=V_0\left[\frac{{(1+t^{\′})}^{1-n}-1}{1-n}\right],$

Sun Chong rising sun formula is:

Q _j＝a·t' ⁱ，

In above-mentioned formula, Q _jrefer to desorption of mash gas cumulative amount; T' refers to the time; V ₀refer to desorption of mash gas speed during t'=0; A, i, n refer to the constant relevant with the gas bearing capacity of coal and structure;

In the normal recovery process in colliery, the size of coal gas outburst amount constantly fluctuates along with passage of time, reaches peak value when driving face (B1) coal breakage starts; And in the normal recovery process in colliery, the coal breakage amount of driving face (B1) is substantially constant, and in the moment that coal breakage starts, the gas emission of driving face (B1) reaches maximum, afterwards along with dilution distinguished and admirable in coal breakage process, gas emission diminishes gradually; Process due to driving face (B1) coal breakage is equivalent to the process of coal body desorb, therefore in the Gas cumulative amount of coal breakage process and coal body desorption process, desorption of mash gas cumulative amount can relative indicatrix each other each other, then the following relational expression of Gas cumulative amount Q' is described:

$Q^{\′}=V_0^{\′}\left[\frac{{(1+t^{\′})}^{1-n}-1}{1-n}\right],$ V ₀' Gas speed when referring to t'=0,

Q'＝a·t' ⁱ。

4., according to the method for the real-time analysis coal and gas prominent danger one of claims 1 to 3 Suo Shu, it is characterized in that: step 1. in, 10m planted agent is distinguished and admirable without branch before and after air velocity transducer (S), nothing is turned round, accessible, section is unchanged;

Leucoaurin sensor (T ₃) for detecting in the air-flow of its position the data of contained methane, these data are as the data of methane contained by driving face air intake.