Gob-side entry retaining Y-shaped air distribution method for gas control

Abstract

The invention belongs to the field of gas control, and discloses a gob-side entry retaining Y-shaped air distribution method for gas control, which comprises the following steps of: investigating mining conditions of a coal seam mining area, collecting data, determining the position of a gas measurement drill hole, and measuring gas parameters of a coal seam and a top plate; calculating the content of gas released by the overlying strata in the initial caving range; calculating the content of gas released by the overlying strata in the normal mining while mining and rising range; calculating the required air volume of the working face as the total air distribution volume by combining the gas content released by the overlying strata, the working face temperature and the number of people; determining the air volume of a main air inlet way and the air volume of an auxiliary air inlet way; and air distribution is conducted on the main air inlet way and the auxiliary air inlet way. According to the method, the gas released by the gob-side entry retaining goaf is considered, the accuracy of air volume calculation is improved, and the air distribution reasonability is improved.

Description

Technical Field

[0001] The invention belongs to the field of gas management, and in particular relates to a "Y" type air distribution method for gob-side tunnel retention used for gas management. Background Art

[0002] During the mining process of gas-containing coal seams, the equilibrium system composed of the original coal-rock mass and gas fluid system is destroyed by the impact of mining. The coal seam roof falls to form a rock-falling area in the goaf, which changes the stress equilibrium state of the coal-rock mass above the goaf. As a result, the original occurrence and flow state of the gas changes, and the gas flows to the free space of the goaf through the formed coal-rock fissure channels. For the coal mining face, the gas gushing locations within the mining area mainly include the upper and lower adjacent coal seams, the surrounding rock of the goaf, the coal left in the goaf, and the goaf of the adjacent mined face. Therefore, it is particularly important to deal with the gas generated by the rock-falling in the goaf and reduce the gas concentration of the return air flow through the air volume ratio. Summary of the invention

[0003] The present invention aims to solve the problem that the gas control methods in the prior art are not effective after the collapse of the goaf in the goaf-retaining lanes, and takes into account the release of gas in the goaf in the goaf-retaining lanes and the release of gas from the coal seams during coal mining, and provides a "Y"-type air distribution method along the goaf-retaining lanes for gas control, so as to reduce the gas concentration in the return air flow as much as possible.

[0004] In order to solve the above technical problems, the technical solution adopted by the present invention is: a "Y" type air distribution method for gob-side tunnel retention for gas control, comprising the following steps:

[0005] Step 1: Investigate the mining conditions in the coal seam mining area, collect data, determine the location of the gas measurement drilling hole, and measure the gas parameters of the coal seam and roof;

[0006] Step 2: Calculate the gas content released from the overlying rock strata within the initial collapse range;

[0007] Step 3: Calculate the gas content released from the overburden strata within the scope of normal mining;

[0008] Step 4: Get the gas content according to steps 2 and 3, and calculate the required air volume of the working face as the total air volume in combination with the working face temperature and number of people;

[0009] Step 5: Determine the actual required air volume of the main air inlet tunnel and the actual required air volume of the auxiliary air inlet tunnel according to the total air volume distribution;

[0010] Step 6: Distribute air to the main air intake channel and the auxiliary air intake channel according to the actual air volume required by the main air intake channel and the auxiliary air intake channel.

[0011] In step 1, the gas parameters measured include: gas content, gas composition, borehole gas flow and attenuation coefficient.

[0012] In step 2, the gas content released from the overlying strata within the initial collapse range is calculated based on the attenuation coefficient, the initial collapse step distance and the length of the working face.

[0013] In step 2, the calculation formula for the gas content released from the overlying strata within the initial collapse range is:

[0014]

[0015] Among them, q1 represents the gas content released by the overburden in the initial collapse range, l represents the length of the working face, h represents the thickness of the immediate roof, a represents the initial collapse step, and ρ 直接顶 represents the direct top density; n represents the gas content in the oil shale, k represents the collapse coefficient, and t1 represents the first collapse time.

[0016] In step 3, the gas content released from the overlying strata within the scope of normal mining during initial collapse is calculated based on the attenuation coefficient, the cutting depth of the coal mining machine and the length of the working face.

[0017] In step 3, the formula for calculating the gas content released from the overburden strata within the scope of normal mining is:

[0018]

[0019] Where: q2 represents the gas content released by the overburden strata within the range of normal mining; l represents the length of the working face, h represents the thickness of the immediate roof, B represents the cutting depth of the coal mining machine, and ρ 直接顶 ——direct top density; n——gas content in oil shale, k represents collapse coefficient, t2 represents the time required to cut one piece of coal.

[0020] In step 4, the calculation formula for the required air volume of the working surface is:

[0021] Q C =max{Q c1 , Q c2 , Q c3};

[0022]

[0023] Q c2 =60×m×v 采 ×S 采 ×K 采高 ×K 采面长 ;

[0024] Q c3 =4N;

[0025] Among them, Q CIndicates the air volume required for the working surface, Q c1 , Q c2 , Q c3 They represent the required air volume according to the gas outburst volume, meteorological conditions and the number of people working at the same time in the mining face. C represents the gas concentration of the return air flow in the working face. q It represents the average absolute gas emission in the return airway of the mining face, q c is the maximum value of the average absolute gas outflow in the air flow of the return air lane of the working face, Indicates the spare air volume coefficient for uneven gas outburst at the mining face, v 采 Indicates the wind speed at the coal mining face, S 采 Represents the average effective cross-sectional area of ​​the coal mining face, K 采高 Indicates the height adjustment coefficient of coal mining face, K 采面长 represents the adjustment coefficient of the length of the coal mining face, t represents the effective ventilation section coefficient, and N represents the number of people working simultaneously on the coal mining face.

[0026] The value of m is 70%.

[0027] The calculation formula for the actual required air volume of the main air inlet tunnel and the actual required air volume of the auxiliary air inlet tunnel is:

[0028] Q z =Q c ÷(1+K)×K;

[0029] Q f =Q f =Q c ÷(1+K)×1;

[0030]

[0031] Where Q z , Q f are respectively the actual required air volume of the main air intake tunnel and the actual required air volume of the auxiliary air intake tunnel, a is the gas outburst ratio of the goaf; b is the coal wall and coal falling outburst ratio; K1 is the gas outburst imbalance coefficient; K2 is the goaf leakage rate; K3 is the ventilation mode coefficient; K is the air distribution ratio, Q C Indicates the air volume required for the working surface.

[0032] The step 4 also includes a step of verifying the required air volume of the working surface, and the verification formula is:

[0033] 0.25S 采 c <4S 采 (m 3 / min);

[0034] In the formula, S 采 Represents the average effective cross-sectional area of ​​the coal mining face.​

[0035] Compared with the prior art, the present invention has the following beneficial effects:

[0036] (1) Compared with the total collapse method, the present invention takes into account the gas released from the goaf along the goaf and a portion of the gas brought out by the air volume, and can more accurately calculate the air volume and distribute the air.

[0037] (2) The present invention considers the gas release amount in the goaf during the initial collapse and the gas emission during mining, thereby improving the accuracy of gas calculation and further improving the rationality of air distribution. BRIEF DESCRIPTION OF THE DRAWINGS

[0038] Figure 1 A schematic diagram of a "Y" type ventilation in an embodiment of the present invention;

[0039] Figure 2 A schematic diagram of a drilling hole for measuring gas parameters in an embodiment of the present invention;

[0040] Figure 3 Schematic diagram of drilling and sealing for gas content testing in an embodiment of the present invention;

[0041] Figure 4 Schematic diagram of gas sample collection in an embodiment of the present invention, wherein (a) is a schematic diagram of natural outflow sampling in a borehole, and (b) is a schematic diagram of sampling with a pump;

[0042] Figure numerals: 1 is a drill hole, 2 is a four-part iron pipe, 3 is Marisan, 4 is a four-part ball valve, 5 is a joint, 6 is an air sample bag, and 7 is an exhaust pipe. DETAILED DESCRIPTION

[0043] In order to make the purpose, technical solutions and advantages of the embodiments of the present invention clearer, the technical solutions in the embodiments of the present invention will be clearly and completely described below. Obviously, the described embodiments are part of the embodiments of the present invention, rather than all the embodiments; based on the embodiments of the present invention, all other embodiments obtained by ordinary technicians in this field without making creative work are within the scope of protection of the present invention.

[0044] The embodiment of the present invention provides a Y-type ventilation and air distribution method for gas control along the gob-retained lane. Figure 1 As shown, it is a schematic diagram of "Y" type ventilation, wherein the track chute and the belt chute are both used as the air inlet lane, and the return air chute is used as the return air lane, forming Y-type ventilation. The embodiment of the present invention includes the following steps:

[0045] Step 1: Investigate the mining conditions of the coal seam mining area, collect data, determine the location of the gas measurement drilling holes, and measure the gas parameters of the coal seam and roof.

[0046] In step 1, the gas parameters measured include: gas content, gas composition, borehole gas flow and attenuation coefficient.

[0047] (1) Determination of gas content and gas composition.

[0048] The construction drilling is used to measure the gas parameters of the coal seam and the immediate roof (gas content, gas gas composition, borehole gas flow and attenuation coefficient). Figure 2 As shown in the figure, holes are drilled along the coal seam on both sides of the transport chute to take samples and measure the gas parameters of the coal seam; holes are drilled on the roof on both sides of the transport chute to take samples and measure the gas parameters of the upper direct roof. Figure 3 As shown, the borehole 1 is sealed with Marisan 3 (polyurethane) and a four-point iron pipe 2, and the borehole gas flow is measured regularly. During the measurement process, gas samples are collected and sent to the laboratory to measure the proportion of each component in the mixed gas. Coal (rock) samples are sampled using core pipes. The design parameters of the borehole for measuring gas parameters in the transport chute are shown in Table 1.

[0049] Table 1 Drilling design parameters for gas parameter determination in transport chute

[0050]

[0051] When drilling underground, coal samples are collected for gas desorption. The gas loss during the sampling process is calculated based on the gas desorption law of the coal samples. The remaining gas is then measured in the laboratory. Finally, the gas content of the coal seam is calculated based on the gas loss, underground gas desorption, remaining gas and coal sample weight. The gas content measured by the degassing method is:

[0052] Connect the four-point iron pipe with the matching joint and put it into the borehole. In order to ensure that the connection of the four-point iron pipe does not leak, the sealing tape should be wrapped around the joint, and the tightness should be appropriate when the pipe clamp is used to tighten the pipe. Then use Marisan (polyurethane) to seal the hole tightly, and the sealing depth should not be less than 6m. After the Marisan expands and solidifies to fix the four-point iron pipe, wait for 0.5 to 1 hour, and then install the ball valve 4 on the four-point iron pipe 2, and make sure the valve is in the closed state. When the gas volume of the borehole is sufficient, you can wait for a certain period of time after sealing the hole and start collecting gas samples; if the gas volume of the borehole is small, collect gas samples on the next day.

[0053] like Figure 4 As shown, during sampling, the gas sample bag 6 and the joint 5 are preferentially used to collect the naturally gushing gas from the four-branch pipe 2. When the gas volume or gas pressure in the borehole is small, a manual sampler (vacuum pump 7) can be used to extract the gas from the four-branch pipe 2 and discharge it into the gas sample bag 6.

[0054] Before collecting gas samples, use the natural gas from the borehole or the gas from the vacuum pump to clean the gas sample bag three times in a row. After the gas sample bag is full, it should be sealed with a spring clamp in time to prevent leakage. After collecting the gas sample, test and analyze the gas components.

[0055] (2) Determination of borehole gas flow and attenuation coefficient.

[0056] After the borehole for gas content measurement is sealed (or the borehole is measured using gas composition), the gas outflow from the borehole is measured using a wet flow meter or a gas meter, and the measurement is done once a day until it becomes basically stable.

[0057] Step 2: Calculate the gas content released from the overlying rock strata within the initial collapse range.

[0058] Specifically, in step 2, the gas content released from the overlying strata within the initial collapse range is calculated based on the attenuation coefficient, the initial collapse step distance and the working face length.

[0059]

[0060] Where: q1——gas content released from the overlying strata within the initial collapse range, m 3 / min;

[0061] l——working surface length, m;

[0062] h——thickness of the immediate roof, m;

[0063] a——first collapse step distance, m;

[0064] ρ 直接顶 ——Direct top density, t / m 3 ;

[0065] n——gas content in oil shale, m 3 / t;

[0066] k——collapse coefficient (determined according to the direct top crushing degree, the smaller the block size, the greater k), 0.4-0.6;

[0067] t1——time to first collapse, min.

[0068] Step 3: Calculate the gas content released from the overlying strata within the scope of normal mining.

[0069] Specifically, in step 3, the gas content released from the overlying strata within the scope of initial cave-in and normal mining is calculated based on the attenuation coefficient, the cutting depth of the coal mining machine and the length of the working face.

[0070]

[0071] Where: q2——gas content released from overburden strata within the range of normal mining, m 3 / min;

[0072] l——working surface length, m;

[0073] h——thickness of the immediate roof, m;

[0074] b——coal mining machine cutting depth, m;

[0075] ρ 直接顶 ——Direct top density, t / m 3 ;

[0076] n——gas content in oil shale, m 3 / t;

[0077] k——collapse coefficient (determined according to the direct top crushing degree, the smaller the block size, the greater k), 0.4-0.6;

[0078] t2——Time required to cut one piece of coal, min.

[0079] Step 4: Based on the gas content obtained in steps 2 and 3, combined with the working face air temperature and number of people, calculate the required air volume of the working face as the total air volume.

[0080] q c =q+max{q1,q2}; (3)

[0081] Where: q ——Average absolute gas outflow in the return airway of the mining face (m 3 / min);

[0082] q c ——The maximum value of gas outflow in the air flow of the return airway of the working face (m 3 / min).

[0083] In step 4, the calculation formula for the required air volume of the working surface is:

[0084] Q C =max{Q c1 , Q c2 , Q c3};(4)

[0085] Where: Q C ——Air volume required for working surface (m 3 / min);

[0086] Q c1 , Q c2 , Q c3 ——According to the required air volume (m3 / min).

[0087]

[0088] Where: C——allowable gas concentration in return air lane and working face (1%);

[0089] ——The spare air volume coefficient for uneven gas outburst at the mining face is obtained by measuring the ratio of the maximum daily absolute gas outburst to the average daily absolute gas outburst over five consecutive days under normal production conditions.

[0090] Q c2 =60×m×v 采 ×S 采 ×K 采高 ×K 采面长 ; (6)

[0091] Where: v 采 ——wind speed at the coal mining face;

[0092] S 采 ——The average effective cross-sectional area of ​​the coal mining face, calculated as the average of the effective cross-sectional area with the maximum and minimum top control distances (m 2 );

[0093] K 采高 ——Coal mining face mining height adjustment coefficient;

[0094] K 采面长 ——coal mining face length adjustment coefficient;

[0095] m——effective ventilation section coefficient, the value is 70%.

[0096] Q c3 =4N; (7)

[0097] Where: N is the number of people working simultaneously on the coal mining face.

[0098] In addition, this embodiment also includes a step of verifying the required air volume of the working surface, and the verification formula is:

[0099] 0.25S 采 c <4S 采 (m 3 / min); (8)

[0100] Where: S 采 ——Average effective cross-sectional area of ​​coal mining face (m 2 ).

[0101] If the working surface requires air volume Q C ​If the above conditions are met, proceed to the next step. If not, adjust the wind speed of the coal mining face.

[0102] Step 5: Allocate the air volume calculated according to the gas outflow volume according to the theoretical calculation formula. The theoretical calculation formula for the air volume of the main air inlet channel is:

[0103]

[0104] On the premise that the total air intake volume is equal to the air volume calculated according to the gas outflow volume, the theoretical calculation formula for the air volume of the auxiliary air intake lane is:

[0105] Q2=(1―K3)K1bq c / C―K2Q1; (10)

[0106] The air distribution ratio is:

[0107]

[0108] Where: Q1, Q2 - theoretically calculated air volume of the main air inlet channel and the air volume of the auxiliary air inlet channel (m 3 / min);

[0109] q c ——The maximum value of gas outflow in the air flow of the return airway of the working face (m 3 / min);

[0110] C——Allowable gas concentration in return air lane and working face (1%);

[0111] a——gas outburst ratio in goaf (40%~60%);

[0112] b——Ratio of coal wall and coal falling (60% to 40%);

[0113] K1——gas outburst imbalance coefficient (1.2~2.1);

[0114] K2——air leakage rate in goaf (10%~30%);

[0115] K3——Ventilation coefficient (0.2 for Y-type ventilation);

[0116] K——Wind ratio.

[0117] Therefore, the actual air volume required for the main air inlet channel and the auxiliary air inlet channel is:

[0118] Q z =Q c ÷(1+K)×K; (12)

[0119] Q f =Q c÷(1+K)×1; (13)

[0120] Where: Q z , Q f ——The actual air volume required for the main air inlet tunnel and the actual air volume required for the auxiliary air inlet tunnel (m 3 / min).

[0121] Step 6: distribute air to the main air inlet lane and the auxiliary air inlet lane according to the actual air volume required by the main air inlet lane and the actual air volume required by the auxiliary air inlet lane. Specifically, the air volume is adjusted by the air door and the air window to achieve air distribution optimization.

[0122] Finally, it should be noted that the above embodiments are only used to illustrate the technical solutions of the present invention, rather than to limit it. Although the present invention has been described in detail with reference to the aforementioned embodiments, those skilled in the art should understand that they can still modify the technical solutions described in the aforementioned embodiments, or replace some or all of the technical features therein with equivalents. However, these modifications or replacements do not cause the essence of the corresponding technical solutions to deviate from the scope of the technical solutions of the embodiments of the present invention.

Claims

1. A "Y" type air distribution method for gob-side tunnel retention for gas control, characterized in that: The following steps are involved: Step 1: Investigate the mining conditions in the coal seam mining area, collect data, determine the location of the gas measurement drilling hole, and measure the gas parameters of the coal seam and roof; Step 2: Calculate the gas content released from the overlying rock strata within the initial collapse range; Step 3: Calculate the gas content released from the overburden strata within the scope of normal mining; Step 4: Based on the gas content obtained in steps 2 and 3, combined with the working face air temperature and number of people, calculate the required air volume of the working face as the total air volume; Step 5: Determine the actual required air volume of the main air inlet tunnel and the actual required air volume of the auxiliary air inlet tunnel according to the total air volume distribution; Step 6: Distribute air to the main air intake channel and the auxiliary air intake channel according to the actual air volume required by the main air intake channel and the auxiliary air intake channel.

2. A "Y" type air distribution method for gob-side tunnel retention for gas control according to claim 1, characterized in that: In step 1, the gas parameters measured include: gas content, gas composition, borehole gas flow and attenuation coefficient.

3. A "Y" type air distribution method for gob-side tunnel retention for gas control according to claim 2, characterized in that: In step 2, the gas content released from the overlying strata within the initial collapse range is calculated based on the attenuation coefficient, the initial collapse step distance and the length of the working face.

4. A "Y" type air distribution method for gob-side tunnel retention for gas control according to claim 3, characterized in that: In step 2, the calculation formula for the gas content released from the overlying strata within the initial collapse range is: Among them, q1 represents the gas content released by the overburden in the initial collapse range, l represents the length of the working face, h represents the thickness of the immediate roof, a represents the initial collapse step, and ρ 直接顶 represents the direct top density; n represents the gas content in the oil shale, k represents the collapse coefficient, and t1 represents the first collapse time.

5. A "Y" type air distribution method for gob-side tunnel retention for gas control according to claim 2, characterized in that: In step 3, the gas content released from the overlying strata within the scope of normal mining during initial collapse is calculated based on the attenuation coefficient, the cutting depth of the coal mining machine and the length of the working face.

6. A "Y" type air distribution method for gob-side entry retention for gas control according to claim 5, characterized in that: In step 3, the formula for calculating the gas content released from the overburden strata within the scope of normal mining is: Where: q2 represents the gas content released by the overburden strata within the range of normal mining; l represents the length of the working face, h represents the thickness of the immediate roof, B represents the cutting depth of the coal mining machine, and ρ 直接顶 represents the direct top density; n represents the gas content in the oil shale, k represents the collapse coefficient, and t2 represents the time required to cut one piece of coal.

7. A "Y" type air distribution method for gob-side entry retention for gas control according to claim 1, characterized in that: In step 4, the calculation formula for the required air volume of the working surface is: Q C =max{Q c1 ,Q c2 ,Q c3 }; Q c2 =60×m×v 采 ×S 采 ×K 采高 ×K 采面长 ; Q c3 =4N; Among them, Q C Indicates the air volume required for the working surface, Q c1 , Q c2 , Q c3 They represent the required air volume according to the gas outburst volume, meteorological conditions and the number of people working at the same time in the mining face, C represents the gas concentration of the return air flow in the working face, q represents the average absolute gas outburst volume in the return airway air flow of the mining face, and q c is the maximum value of the average absolute gas outflow in the air flow of the return air lane of the working face, Indicates the spare air volume coefficient for uneven gas outburst at the mining face, v 采 Indicates the wind speed at the coal mining face, S 采 Represents the average effective cross-sectional area of ​​the coal mining face, K 采高 Indicates the height adjustment coefficient of coal mining face, K 采面长 represents the adjustment coefficient of the length of the coal mining face, t represents the effective ventilation section coefficient, and N represents the number of people working simultaneously on the coal mining face.

8. A "Y" type air distribution method for gob-side entry retention for gas control according to claim 7, characterized in that: The value of m is 70%.

9. A "Y" type air distribution method for gob-side entry retention for gas control according to claim 1, characterized in that: The calculation formula for the actual required air volume of the main air inlet tunnel and the actual required air volume of the auxiliary air inlet tunnel is: Q z =Q c ÷(1+K)×K; Q f =Q f =Q c ÷(1+K)×1; Where Q z , Q f are respectively the actual required air volume of the main air intake tunnel and the actual required air volume of the auxiliary air intake tunnel, a is the gas outburst ratio of the goaf; b is the coal wall and coal falling outburst ratio; K1 is the gas outburst imbalance coefficient; K2 is the goaf leakage rate; K3 is the ventilation mode coefficient; K is the air distribution ratio, Q C Indicates the air volume required for the working surface.

10. A "Y" type air distribution method for gob-side entry retention for gas control according to claim 1, characterized in that: The step 4 also includes a step of verifying the required air volume of the working surface, and the verification formula is: 0.25S 采 <Q c <4S 采 (m 3 / min); In the formula, S 采 Represents the average effective cross-sectional area of ​​the coal mining face.

Images