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Past Questions
Which of the following is known as point load index?
A)
(14+0.175 D)Is
B)
P/D2
C)
σ = 2F / DL
D)
S = C + s tan F
The Point Load Test (PLT) is a widely used method in rock mechanics to estimate the uniaxial compressive strength (UCS) of rock samples. The Point Load Index (Is) is defined as the load at failure (P) divided by the square of the equivalent core diameter (D). The formula is expressed as: Is = P / D², where P is the failure load and D is the distance between the loading points (for a standard cylindrical sample).
Analysis of other options:
- Option (a) represents a correlation formula used to estimate the Uniaxial Compressive Strength (UCS) from the point load index, not the index itself.
- Option (c) resembles a formula for calculating stress (σ) in a different mechanical context (likely a bearing or shear stress formula).
- Option (d) is the Mohr-Coulomb failure criterion (S = C + σ tan φ), which describes the shear strength of a material based on cohesion (C) and the angle of internal friction (φ).
The point beyond which the mineral/coal cannot be economically extracted is called
A)
Stripping ratio
B)
Cost of stripping ratio
C)
Break-even ratio
D)
Limit of extraction
In open-pit mining, the economic viability of extracting ore is determined by the relationship between the cost of removing overburden (waste) and the value of the mineral recovered.
1. **Break-even ratio (or Break-even Stripping Ratio - BESR)**: This is the specific ratio of waste to ore at which the cost of extraction equals the value of the mineral. Beyond this point, the cost of stripping the overburden exceeds the revenue generated from the mineral, making further extraction economically unfeasible. Therefore, it defines the ultimate limit of the open pit.
2. **Why other options are incorrect**:
- **Stripping ratio (a)**: This is a general term referring to the ratio of overburden volume to the volume of ore. It is a measurement, not necessarily the 'point' of economic limit.
- **Cost of stripping ratio (b)**: This is not a standard technical term; it refers to the expense associated with the stripping ratio rather than the limit of extraction.
- **Limit of extraction (d)**: While descriptive, this is a general term. In mining engineering and economics, the precise technical term for the economic threshold is the 'Break-even ratio'.
In transverse stoping methods, stoping operation advances:
A)
Along strike direction
B)
Along oblique direction
C)
From hangwall to footwall
D)
Any of the above
In underground mining, stoping methods are generally classified based on the direction of advance relative to the orebody's geometry.
1. **Longitudinal Stoping**: The stoping operation advances along the strike of the orebody. This is typically used for narrow veins.
2. **Transverse Stoping**: The stoping operation advances perpendicular to the strike, moving from the hanging wall (the rock above the orebody) toward the footwall (the rock below the orebody), or vice versa. This method is typically employed in wide orebodies where the width of the ore exceeds the stable span of the unsupported roof.
Therefore, in transverse stoping, the advance is across the width of the orebody (from hangwall to footwall), making option 'c' the correct answer. Option 'a' describes longitudinal stoping, and option 'b' is not a standard classification for these primary stoping methods.
Match the following: Class of fire: Fires of A. Class A: 1. Inflammable gases B. Class B: 2. Inflammable liquids C. Class C: 3. Combustible solid materials D. Class D: 4. Metals
A)
A-2, B-3, C- 1, D-4
B)
A-3, B-1, C- 2, D-4
C)
A-4, B-2, C- 1, D-3
D)
A-3, B-2, C- 1, D-4
The classification of fires is standardized to determine the correct extinguishing agent. Based on the standard fire classification used in industrial and mine safety contexts:
1. Class A fires involve ordinary combustible solid materials such as wood, paper, cloth, and rubber. (A matches with 3)
2. Class B fires involve flammable liquids such as petrol, diesel, oils, and solvents. (B matches with 2)
3. Class C fires involve flammable gases such as methane, propane, and butane. (C matches with 1)
4. Class D fires involve combustible metals such as magnesium, titanium, and sodium. (D matches with 4)
Matching these pairs gives: A-3, B-2, C-1, D-4. This corresponds exactly to option (d).
Water gas is formed by passing ......... over heated ............
A)
Air, water
B)
Carbon monoxide, surface
C)
Water, iron
D)
Steam, hydrocarbons
Water gas (also known as synthesis gas or syngas) is a mixture of carbon monoxide (CO) and hydrogen (H2). It is produced industrially by passing steam (water vapor) over the surface of incandescent (heated) carbon or hydrocarbons (such as coke or coal). The chemical reaction is: C + H2O → CO + H2.
Analysis of options:
- Option (a) Air and water: Passing air over water does not produce water gas.
- Option (b) Carbon monoxide and surface: This describes components/locations, not the reactants required to form the gas.
- Option (c) Water and iron: While iron can be a catalyst in some reactions, the primary fuel source for water gas must be a carbon-based material (hydrocarbon).
- Option (d) Steam and hydrocarbons: This is the correct chemical process. Steam (H2O) reacts with heated carbon/hydrocarbons (C) to produce the CO and H2 mixture.
Match the following with respect to Graham’s ratio: Status of fire: Graham’s Ratio A. Active fire: 1. 2.0 B. Necessity for checkup: 2. 1.0 C. Existence for fire: 3. 0.4 D. Spontaneous heating approaches active fire: 4. 0.5 E. Normal: 5. 3.0
A)
A-1, B-4, C-5, D-2, E-3
B)
A-4, B-5, C-2, D-1, E-3
C)
A-5, B-4, C-2, D-1, E-3
D)
A-5, B-4, C-2, D-3, E-1
Graham's Ratio is a critical tool used in mine safety to detect and monitor spontaneous combustion in coal mines. It is calculated as the ratio of Carbon Monoxide (CO) to Oxygen deficiency (O2 deficiency = 20.9% - actual O2%).
According to established mining safety standards for Graham's Ratio:
1. Normal/No heating: Ratio is typically very low (around 0.4 or less).
2. Existence of fire/Heating: Ratio starts to rise, typically around 1.0.
3. Spontaneous heating approaching active fire: Ratio increases further, often around 2.0.
4. Necessity for checkup/Warning: Ratio is in the intermediate range, often around 3.0 or 4.0 depending on the specific scale used, but in the context of this specific matching set, the values are mapped as follows:
- Normal (E) matches with 0.4 (3).
- Existence of fire (C) matches with 1.0 (2).
- Spontaneous heating approaches active fire (D) matches with 2.0 (1).
- Necessity for checkup (B) matches with 0.5 (4) - Note: In some specific regional scales, lower ratios indicate early warning, but based on the options provided, B-4 is the only fit for the remaining logic.
- Active fire (A) matches with 3.0 (5).
Matching the pairs:
A (Active fire) -> 5 (3.0)
B (Necessity for checkup) -> 4 (0.5)
C (Existence for fire) -> 2 (1.0)
D (Spontaneous heating approaches active fire) -> 1 (2.0)
E (Normal) -> 3 (0.4)
This corresponds to Option c: A-5, B-4, C-2, D-1, E-3.
The factor of safety of a slope will be less than one, when (consider that the cohesion is zero):
A)
Slope angle is less than the friction angle
B)
Slope angle is greater than friction angle
C)
Slope angle is equal to the friction angle
D)
None of the above
In soil mechanics and slope stability analysis, the Factor of Safety (FoS) is defined as the ratio of the resisting forces (shear strength) to the driving forces (shear stress). For a cohesionless soil (where cohesion c = 0), the shear strength is derived solely from internal friction. The formula for the Factor of Safety of an infinite slope in cohesionless soil is: FoS = tan(φ) / tan(β), where φ is the angle of internal friction and β is the slope angle.
1. If β < φ (Slope angle is less than friction angle): tan(φ) > tan(β), therefore FoS > 1. The slope is stable.
2. If β = φ (Slope angle is equal to friction angle): tan(φ) = tan(β), therefore FoS = 1. The slope is in a state of limiting equilibrium.
3. If β > φ (Slope angle is greater than friction angle): tan(φ) < tan(β), therefore FoS < 1. The driving force exceeds the resisting force, and the slope will fail.
Therefore, the factor of safety is less than one when the slope angle is greater than the friction angle.
The factor of safety of a slope is 1.74 for the slope angle of 45°. For which slope angle, the factor of safety will be equal to 1.0 (consider that the cohesion is zero):
A)
~50°
B)
~55°
C)
~60°
D)
~65°
Step-by-step derivation:
1. **Identify the Formula**: For a slope where cohesion (c) is zero, the Factor of Safety (FoS) is defined by the ratio of resisting forces to driving forces. For a simple planar failure (or infinite slope), the formula is:
FoS = (tan φ) / (tan θ)
Where:
- φ = angle of internal friction of the material
- θ = slope angle
2. **Use Given Data to find φ**:
Given: FoS = 1.74 when θ = 45°
1.74 = tan φ / tan(45°)
Since tan(45°) = 1:
tan φ = 1.74
3. **Calculate the Critical Slope Angle (θ_crit)**:
We need to find the angle θ for which the Factor of Safety (FoS) equals 1.0.
1.0 = tan φ / tan θ_crit
tan θ_crit = tan φ
From step 2, we know tan φ = 1.74.
Therefore, tan θ_crit = 1.74
4. **Solve for θ_crit**:
θ_crit = arctan(1.74)
Using a calculator: θ_crit ≈ 60.11°
5. **Conclusion**: The calculated angle is approximately 60°, which matches option (c).
On what factors does the specific gravity of coal depend?
A)
Its calorific value and ash content
B)
Its ignition temperature and calorific value
C)
Type of coal and ash content
D)
Type of coal and calorific value
The specific gravity (SG) of coal is a measure of its density relative to water and is primarily influenced by two main factors: 1. Type of Coal (Rank): As coal matures from peat to lignite, bituminous, and finally anthracite, the carbon content increases and the volatile matter decreases. Higher-rank coals (like anthracite) generally have a higher specific gravity than lower-rank coals (like lignite) because they are more compact and carbon-dense. 2. Ash Content: Ash consists of inorganic minerals (such as silica, alumina, and iron oxides) which are significantly denser than the organic carbonaceous matter of the coal. Therefore, an increase in the percentage of ash content directly increases the overall specific gravity of the coal sample.
Why other options are incorrect:
- Calorific value (Options a, b, d) is a result of the coal's composition (carbon content) rather than a factor that determines its physical density/specific gravity.
- Ignition temperature (Option b) is a thermal property related to combustion and does not influence the physical density of the material.
With the increase in depth of cover, surface subsidence value:
A)
Increases
B)
decreases
C)
is constant
D)
does not show any trend
In mining engineering and geomechanics, surface subsidence is the downward movement of the earth's surface resulting from the removal of subsurface materials (such as coal or minerals). The relationship between the depth of cover (overburden) and the maximum subsidence value is inversely proportional. As the depth of cover increases, the overlying rock strata act as a more substantial beam or arch, distributing the stress more widely and absorbing more of the deformation within the strata themselves before it reaches the surface. Consequently, for a given extraction thickness and area, the percentage of subsidence reaching the surface decreases as the depth of the mine increases. Conversely, shallow mines typically experience more severe and concentrated surface subsidence.
For circular slope failure the factor of safety can be expresses as
A)
Tensile strength available to resist sliding / Tensile strength induces sliding
B)
Shear strength available to resist sliding / Shear strength induces sliding
C)
Shear strength induces sliding / Shear strength available to resist sliding
D)
None of these
In geotechnical engineering and slope stability analysis, the Factor of Safety (FoS) is defined as the ratio of the resisting forces (or strengths) to the driving forces (or stresses) that cause failure. For a circular slope failure, the failure occurs when the shear stress induced along the slip surface exceeds the available shear strength of the soil.
1. **Resisting Force:** This is the shear strength of the soil along the failure surface, typically calculated using the Mohr-Coulomb failure criterion (c + σ tan φ).
2. **Driving Force:** This is the shear stress induced by the weight of the soil mass and other external loads, which acts to push the soil mass downward along the circular arc.
Therefore, Factor of Safety (FoS) = (Total Resisting Shear Strength) / (Total Driving Shear Stress).
- Option (a) is incorrect because slope stability is governed by shear failure, not tensile failure.
- Option (c) is incorrect because it represents the reciprocal of the factor of safety (Driving/Resisting), which would indicate instability if the value is greater than 1.
- Option (b) correctly identifies the ratio of available strength to the strength induced (demanded) by the sliding mechanism.
If the full/maximum subsidence is occurred more than one point, then the subsidence is called:
A)
Critical subsidence
B)
Subcritical subsidence
C)
Supercritical subsidence
D)
Trough subsidence
In mining engineering, subsidence is categorized based on the width of the extraction area relative to the depth of cover.
1. **Subcritical Subsidence**: Occurs when the width of the extracted area is small relative to the depth. The maximum possible subsidence is not reached, and the subsidence profile is asymmetrical, with the maximum subsidence occurring closer to the edge of the panel.
2. **Critical Subsidence**: This is the threshold state where the width of the extraction is just sufficient for the maximum possible subsidence to occur at a single central point.
3. **Supercritical Subsidence**: Occurs when the width of the extraction area exceeds the critical width. In this state, the maximum possible subsidence is achieved over a wide area (a 'trough'), meaning the full/maximum subsidence occurs at more than one point across the center of the panel.
4. **Trough Subsidence**: This is a general term describing the shape of the surface depression, but 'Supercritical' specifically defines the state where the maximum subsidence value is maintained across multiple points.
The pore water pressure ............ the slope stability:
A)
Increases
B)
Decreases
C)
Does not have any effect
D)
May increase or decrease
Slope stability is primarily governed by the shear strength of the soil or rock mass. According to the Mohr-Coulomb failure criterion, the shear strength (τ) is defined as: τ = c + (σ - u) tan(φ), where 'c' is cohesion, 'σ' is the total normal stress, 'u' is the pore water pressure, and 'φ' is the angle of internal friction. The term (σ - u) represents the 'effective stress' (σ'). As pore water pressure (u) increases, the effective stress (σ') decreases. Since the frictional component of shear strength is directly proportional to the effective stress, an increase in pore water pressure reduces the overall shear strength of the material. When the shear strength decreases while the driving forces (gravity/load) remain constant, the Factor of Safety (FoS) decreases, thereby decreasing the stability of the slope. Therefore, pore water pressure acts as a destabilizing force.
One Murg is equal to.........Ns2/m8.
A)
0.0981
B)
0.00981
C)
9.81
D)
None of the above
The 'Murg' is a non-SI unit used in rock mechanics and mining engineering to describe the 'blastability' or 'rock mass quality' index, specifically associated with the work of some researchers in the field of rock fragmentation. In the context of these specific units, 1 Murg is defined as 0.00981 Ns²/m⁸. This unit is derived from the relationship between the energy required to fragment a unit volume of rock and the physical properties of the rock mass. Specifically, it relates to the gravitational constant (g ≈ 9.81 m/s²) scaled by a factor of 10⁻³ (1/1000), resulting in 0.00981. Therefore, option (b) is the correct numerical value.
A coal pillar of 36 m x 36 m (centre to centre) is situated at a depth of 150 m. The width of gallery is 6.0 m. Considering the unit weight of rock 0.025 MPa/m, the load on the pillar, calculated by tributary area method would be:
A)
5.4 MPa
B)
5.9 MPa
C)
5.8 MPa
D)
5.1 MPa
Tributary area=36*36=1296, pillar=30*30=900 (36-6), sigma= gamma*H*trib/pillar =0.025*150*1296/900=5.4 MPa.
Anisotropy is defined as:
A)
Rock mass does not have same properties in all the directions
B)
Rock mass have same properties in all the directions
C)
Rock mass does not have same properties in one direction
D)
Rock mass have same properties in one direction
In rock mechanics and materials science, 'Anisotropy' refers to the characteristic of a material where its physical and mechanical properties (such as strength, elasticity, or permeability) vary depending on the direction in which they are measured. This is common in sedimentary rocks (like shale) due to bedding planes or metamorphic rocks (like schist) due to foliation.
- Option (a) is correct because it accurately describes the directional dependence of properties.
- Option (b) describes 'Isotropy', which is the opposite of anisotropy (where properties are identical in all directions).
- Options (c) and (d) are logically flawed as they refer to 'one direction', whereas anisotropy is a comparison of properties across multiple different directions.
As per Unal, suggested bolt length should be:
A)
L = (B * (100 - RMR)) / 200
B)
L = B / 2
C)
L = B / 3
D)
L = 2/3 B
In rock mechanics and underground excavation support design, various empirical formulas are used to determine the required length of rock bolts. Unal (1983) proposed an empirical relationship to estimate the support pressure and the length of the rock bolts based on the Rock Mass Rating (RMR) system developed by Bieniawski. According to Unal's suggestion, the length of the bolt (L) is calculated as a function of the span or width of the excavation (B) and the RMR value. The specific formula is L = [B * (100 - RMR)] / 200. This formula implies that as the rock quality (RMR) decreases, the required bolt length increases to anchor the unstable rock mass into the more stable rock zone. Options b, c, and d represent generic rules of thumb (such as the common rule that bolt length should be roughly half the span), but they are not the specific empirical formula attributed to Unal.
Coal mines roof rating (CMRR) was given by:
A)
Z T Bieniawski
B)
Mark and Molinda
C)
Marinos and Hoek
D)
Laubscher and Taylor
The Coal Mine Roof Rating (CMRR) is a widely used empirical system specifically designed to evaluate the stability of coal mine roofs to determine the appropriate support requirements. It was developed by Mark and Molinda (specifically B.G. Mark and others at the National Institute for Occupational Safety and Health - NIOSH). The system uses parameters such as rock strength, bedding plane spacing, and moisture to assign a numerical rating to the roof rock.
Analysis of other options:
- Option (a) Z.T. Bieniawski developed the Rock Mass Rating (RMR) system, which is a general rock mass classification system used across various types of mining and civil engineering, not specifically for coal mine roofs.
- Option (c) Marinos and Hoek are known for their work on the Geological Strength Index (GSI), which focuses on the structure and surface condition of rock masses.
- Option (d) Laubscher developed the Mining Rock Mass Rating (MRMR) system, which is an adaptation of RMR specifically for caving and large-scale mining operations.
Development works are going on with 35% recovery from a coal seam. The pillars developed during development stage are having strength of 13.27 MPa with factor of safety 2.3. Considering the unit weight of rock 0.025 MPa/m, the depth of cover of the coal seam is:
A)
~200 m
B)
~250 m
C)
~150 m
D)
~100 m
FoS-based stress=13.27/2.3=5.77 MPa = gamma H/(1-e). With e=0.35 => H=5.77*0.65/0.025~150m.
If the porosity in rock mass is higher-
A)
Density will be also higher
B)
Modulus of elasticity will be lower
C)
Transmission of the sound waves will be higher
D)
Internal friction and rock stability will be higher
Porosity refers to the ratio of the volume of voids to the total volume of the rock mass. Here is the step-by-step analysis of the options:
1. **Density (Option a):** Density is mass per unit volume. Since voids (pores) are filled with air or water (which are much less dense than rock minerals), an increase in porosity leads to a decrease in the overall bulk density of the rock. Therefore, option (a) is incorrect.
2. **Modulus of Elasticity (Option b):** The Modulus of Elasticity (Young's Modulus) measures the stiffness of a material. Voids act as discontinuities and points of weakness within the rock matrix. As porosity increases, the effective cross-sectional area of the solid mineral skeleton decreases, reducing the rock's ability to resist deformation. Consequently, the rock becomes less stiff, and the modulus of elasticity decreases. This is a fundamental principle in rock mechanics. Therefore, option (b) is correct.
3. **Transmission of Sound Waves (Option c):** Sound waves (seismic waves) travel more efficiently through dense, continuous solid media. Pores act as scattering centers and barriers that attenuate the signal and slow down the wave velocity. Higher porosity generally leads to lower sound wave velocity and transmission. Therefore, option (c) is incorrect.
4. **Internal Friction and Stability (Option d):** Higher porosity generally weakens the structural integrity of the rock. It increases the likelihood of collapse under load and typically reduces the overall shear strength and stability of the rock mass. Therefore, option (d) is incorrect.
In which method of mining, slot raise are not required:
A)
VCR method
B)
Sublevel method
C)
Room and pillar method
D)
Cut and fill method
A 'slot raise' is a vertical or inclined opening created to provide a free face for blasting in underground mining, typically used in stoping methods to initiate the caving or slicing of the ore body.
1. **VCR (Vertical Crater Retreat) method**: This method relies heavily on the creation of a slot (often using large diameter holes) to create the initial free face for the crater blasting process.
2. **Sublevel method**: This is a bulk mining method where slot raises are essential to create the initial void (slot) from which the rest of the sublevel is blasted in slices.
3. **Cut and fill method**: In many variations of stoping (including cut and fill), slotting is used to initiate the extraction of a block of ore.
4. **Room and pillar method**: This is a horizontal mining method used for flat-lying deposits. It involves excavating 'rooms' while leaving 'pillars' of ore for support. The excavation is done horizontally using drilling and blasting against the face of the room; it does not involve vertical stoping or the need for a vertical slot raise to initiate the collapse or retreat of a block. Therefore, slot raises are not required in this method.
The incubation period in underground coal mine is used for deciding:
A)
Size of coal pillar
B)
Number of pillars in depillaring panel/size of depillaring panel
C)
Width of the gallery
D)
Diagonal line of extraction
In underground coal mining, specifically during depillaring operations, the 'incubation period' refers to the time interval between the initiation of a stress-induced change (such as the start of pillar extraction) and the actual occurrence of a spontaneous combustion event or a major strata failure (like a bump or cave-in).
1. Why Option (b) is correct: The incubation period is critical for planning the 'size of the depillaring panel' and the 'number of pillars' to be extracted. If the incubation period is short, the panel size must be kept small to ensure that the coal is extracted and the area is sealed or goafed before spontaneous combustion can ignite. If the panel is too large, the time taken to extract all pillars may exceed the incubation period, leading to hazardous fires.
2. Why other options are incorrect:
- Option (a) Size of coal pillar: Pillar size is determined by the depth of cover, strength of coal, and safety factors (e.g., using formulas like CMR guidelines), not by the incubation period.
- Option (c) Width of the gallery: Gallery width is decided based on the equipment used and the stability of the roof/sides (geotechnical properties).
- Option (d) Diagonal line of extraction: The line of extraction is decided based on the direction of the main dip, ventilation requirements, and the sequence of extraction to manage stress, rather than the chemical/thermal incubation period.
While determining the tri-axial compressive strength in the laboratory the lateral pressure acting of the rock sample known as:
A)
Atmospheric pressure
B)
Hydrostatic pressure
C)
Deviatric pressure
D)
None of these
In a tri-axial compressive strength test, a cylindrical rock sample is subjected to a constant lateral pressure (confining pressure) while an increasing axial load is applied until failure. This lateral pressure is typically applied using a fluid (oil or water) within a pressure chamber, which exerts an equal pressure in all lateral directions. This type of uniform pressure is known as hydrostatic pressure (or confining pressure).
Analysis of other options:
- Atmospheric pressure (a) is the pressure exerted by the weight of the atmosphere and is not the controlled lateral pressure used to simulate depth in rock mechanics tests.
- Deviatoric pressure (c), or deviatoric stress, is the difference between the axial stress (σ1) and the lateral confining stress (σ3). It is the 'extra' stress applied along the axis, not the lateral pressure itself.
- Therefore, the lateral pressure acting on the sample is the hydrostatic/confining pressure.
Which of the following excavator is preferred to be used for a bench height of 30 m or more?
A)
Bucket wheel Excavator
B)
Shovel
C)
Front and end loader
D)
Dragline
In open-cast mining, the choice of excavator depends heavily on the bench height and the nature of the material.
1. **Dragline (Option d):** Draglines are designed for very deep excavations and high bench heights. Because they cast their bucket and pull it back toward the machine using a cable system, they can reach depths and heights far exceeding those of shovels or loaders. They are typically used for overburden removal where bench heights can reach 30m to 60m or more.
2. **Shovel (Option b):** Electric or hydraulic shovels are highly productive but are limited by the reach of their boom and the stability of the face. They are generally preferred for bench heights ranging from 10m to 20m. A 30m bench would be unstable and unsafe for a standard shovel to operate from the floor.
3. **Bucket Wheel Excavator (Option a):** BWEs are used for continuous mining of soft materials. While they can be massive, they are typically used for specific terrace-mining methods rather than general high-bench excavation in the same manner as a dragline.
4. **Front and End Loader (Option c):** Loaders are used for loading material into haul trucks from a flat surface or low face. They have very limited vertical reach and are unsuitable for high benches.
Therefore, for a bench height of 30m or more, the Dragline is the preferred equipment due to its superior reach and digging depth.
Which is associated only with longwall mining?
A)
Hydraulic prop
B)
Powered support
C)
Shuttle car
D)
Roadheader
Step-by-step reasoning:
1. **Analyze the Question**: The question asks for the equipment associated *only* with longwall mining, meaning it is not typically used in other methods like Bord and Pillar or Room and Pillar mining.
2. **Evaluate Option (a) Hydraulic prop**: Hydraulic props are versatile support systems used in various mining methods, including Bord and Pillar, to support the roof temporarily or permanently.
3. **Evaluate Option (b) Powered support**: Also known as 'Shield Supports', these are massive, automated hydraulic supports specifically designed for longwall mining. They support the roof of the longwall face and advance automatically as the shearer moves forward, allowing the roof behind them to collapse (the goaf). They are not used in Bord and Pillar mining.
4. **Evaluate Option (c) Shuttle car**: Shuttle cars are primarily used in Bord and Pillar or Room and Pillar mining to transport coal from the working face to the feeder breaker or conveyor belt.
5. **Evaluate Option (d) Roadheader**: A roadheader is a versatile cutting machine used for developing galleries, tunnels, or in room and pillar mining. While it can be used in some longwall setups for development, it is not exclusive to longwall mining.
6. **Conclusion**: Powered supports (shields) are the defining characteristic of modern longwall mining systems.