VENTILATION AND COOLING IN UNDERGROUND MINES     

 

M.J. Howes

 

 

The main objective of mine ventilation is the provision of sufficient quantities of air to all the working places and travel ways in an underground mine to dilute to an acceptable level those contaminants which cannot be controlled by any other means. Where depth and rock temperatures are such that air temperatures are excessive, mechanical refrigeration systems may be used to supplement the beneficial effects of ventilation.

 

The Mine Atmosphere

The composition of the gaseous envelope encircling the earth varies by less than 0.01% from place to place and the constitution of “dry” air is usually taken as 78.09% nitrogen, 20.95% oxygen, 0.93% argon and 0.03% carbon dioxide. Water vapour is also present in varying amounts depending on the air temperature and pressure and the availability of free water surfaces. As ventilation air flows through a mine, the concentration of water vapour may change significantly and this variation is the subject of the separate study of psychrometry. To define the state of a water vapour and dry air mixture at a particular point requires the three measurable independent properties of barometric pressure, dry bulb and wet bulb temperatures.

 

Ventilation Requirements

The contaminants to be controlled by dilution ventilation are primarily gases and dust, although ionizing radiations associated with naturally occurring radon may present problems, especially in uranium mines and where the background uranium concentrations of the host or adjacent rocks are elevated. The amount of air required for dilution control will depend on both the strength of the contaminant source and the effectiveness of other control measures such as water for dust suppression or methane drainage systems in coal mines. The minimum dilution air flow rate is determined by the contaminant requiring the greatest dilution quantity with due cognizance of the possible additive effects of mixtures and synergism where one contaminant can increase the effect of another. Overriding this value could be a minimum air velocity requirement which is typically 0.25 m/s and increasing as air temperatures also increase.

 

Diesel-powered equipment ventilation

In mechanized mines using diesel-powered mobile equipment and in the absence of continuous gas monitoring, exhaust gas dilution is used to determine the minimum ventilation air requirements where they operate. The amount of air required normally ranges between 0.03 and  per kW of rated power at the point of operation depending on the type of the engine and whether any exhaust gas conditioning is being used. Continuing developments in both fuel and engine technology are providing lower engine emissions while catalytic converters, wet scrubbers and ceramic filters may further reduce the leaving concentrations of carbon monoxide/aldehydes, oxides of nitrogen and diesel particulates respectively. This helps in meeting increasingly stringent contaminant limits without significantly increasing exhaust dilution rates. The minimum possible dilution limit of  per kW is determined by the carbon dioxide emissions which are proportional to engine power and unaffected by exhaust gas conditioning.

 

Diesel engines are about one-third efficient at converting the energy available in the fuel to useful power and most of this is then used to overcome friction resulting in a heat output which is about three times the power output. Even when hauling rock up a decline in a truck, the useful work done is only about 10% of energy available in the fuel. Higher diesel engine powers are used in larger mobile equipment which require bigger excavations to operate safely. Allowing for normal vehicle clearances and a typical diesel exhaust gas dilution rate of  per kW, the minimum air velocities where diesels operate average about 0.5 m/s.

 

Ventilation of different mining methods

Although the setting of general air quantity requirements is not appropriate where detailed mine and ventilation planning information is available or possible, they are supportive of the criteria being used for design. Deviations from normal values generally can be explained and justified, for instance, in mines with heat or radon problems. The general relationship is:

 

 

where t is the annual production rate in million tonnes per annum (Mtpa),  is a variable air quantity factor which is directly related to production rate and  is the constant air quantity required to ventilate the mine infrastructure such as the ore handling system. Typical values of  are given in table 74.1.

 

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Table 74.1.    Design air quantity factors

 

Mining method  

Block-caving

50

Room-and-pillar (Potash)

75

Sub-level caving

120

Open stoping

large >.5 Mtpa

small <.5 Mtpa

 

160

240

Mechanized cut-and-fill

320

Non-mechanized mining

400

 

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The constant air quantity  is mainly dependent on the ore handling system and, to a certain extent, on the overall mine production rate. For mines where rock is transported through a decline using diesel powered truck haulage or there is no crushing of the mined rock, a suitable value of  is . This typically increases to  when using underground crushers and skip hoisting with underground maintenance areas. As the ore handling system become more extensive (i.e., using conveyors or other ore transfer systems),  can further increase by up to 50%. On very large mines where multiple shaft systems are used, the constant air quantity  is also a multiple of the number of shaft systems required.

 

Cooling Requirements

 

Design thermal conditions

The provision of suitable thermal conditions to minimize the dangers and adverse effects of heat stress may require mechanical cooling in addition to the ventilation necessary to control contaminants. Although the applied heat stress is a complex function of climatic variables and physiological responses to them, in practical mining terms it is the air velocity and wet bulb temperature that have the greatest influence. This is illustrated by the clothing-corrected air cooling powers  given in table 74.2. Underground the radiant temperature is taken to be equal to the dry bulb temperature and 10  higher than the wet bulb temperature. The barometric pressure and the clothing regime are typical for underground work (i.e., 110 kPa and 0.52 clothing units).

 

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Table 74.2.     Clothing-corrected air cooling powers

 

Air velocity (m/s)

Wet bulb temperature (°C)

 

20.0

22.5

25.0

27.5

30.0

32.5

0.1

176

153

128

100

70

37

0.25

238

210

179

145

107

64

0.5

284

254

220

181

137

87

1.0

321

290

254

212

163

104

 

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An air velocity of 0.1 m/s reflects the effect of natural convection (i.e., no perceivable airflow at all). An air velocity of 0.25 m/s is the minimum normally allowed in mining and 0.5 m/s would be required where the wet bulb temperature exceeds 25. With respect to achieving thermal equilibrium, the metabolic heat resulting from typical work rates are: rest, ; light work, 115 to , medium work, 150 to ; and hard work, 200 to . Design conditions for a specific mine application would be determined from a detailed optimization study. Generally, optimum wet bulb temperatures are between 27.5  and 28.5  with the lower temperatures applicable to less mechanized operations. Work performance decreases and the risk of heat-related illness increases significantly when the wet bulb temperature exceeds 30.0 , and work should not normally continue when the wet bulb temperature is greater than 32.5 .

 

Mine heat loads

The mine refrigeration load is the mine heat load less the cooling capacity of the ventilation air. The mine heat load includes the effects of auto-compression of the air in the intake airways (the conversion of potential energy to enthalpy as the air flows down into the mine), heat flow into the mine from the surrounding rock, heat removed from the rock broken or any fissure water before they are removed from the intakes or working sections of the mine, and the heat resulting from the operation of any equipment used in the ore breaking and transportation processes. The cooling capacity of the ventilation air depends on both the design thermal environmental conditions in the working places and the actual climatic conditions on surface.

 

Although the relative contributions of each heat source to the total is site specific, auto-compression is usually the main contributor at between 35 and 50% of the total. As the depth of mining increases, auto-compression can cause the cooling capacity of the air to become negative and the effect of supplying more air is to increase the mine refrigeration load. In this case, the amount of ventilation supplied should be the minimum consistent with meeting contaminant control and increasing amounts of refrigeration are required to provide productive and safe working conditions. The depth of mining at which refrigeration becomes necessary will depend primarily on the surface climatic conditions, the distance the air travels through the intake airways before it is used and the extent to which large equipment (diesel or electric powered) is used.

 

Primary Ventilation Systems

 

Networks

Primary ventilation systems or networks are concerned with ensuring the flow of air through interconnected mine openings. The overall ventilation network has junctions where three or more airways meet, branches that are airways between junctions and meshes which are closed paths traversed through the network. Although most mine ventilation networks are ramified with hundreds or even thousands of branches, the number of main intake (branch between surface and the mine workings) and return or exhaust (branch between the workings and surface) airways is usually limited to less than ten.

 

With large numbers of branches in a network, determining a flow pattern and establishing the overall pressure loss is not straightforward. Although many are in simple series or parallel arrangement which can be solved algebraically and precisely, there will be some compound sections requiring iterative methods with convergence to an acceptable tolerance. Analogue computers have been successfully used for network analysis; however, these have been superseded by less time-consuming digital methods based on the Hardy Cross approximation technique developed to solve water flow networks.

 

Airway resistance and shock losses

The resistance to airflow of a tunnel or mine opening is a function of its size and surface roughness and the resultant pressure loss depends on this resistance and the square of the air velocity. By adding energy to the system, a pressure can be generated which then overcomes the pressure loss. This may occur naturally where the energy is provided by heat from the rock and other sources (natural ventilation). Although this used to be the main method of providing ventilation, only 2 to 3% of the energy is converted and, during hot summers, the rock may actually cool the intake air resulting in flow reversals. In modern mines a fan is normally used to provide energy to the air stream which then overcomes the pressure loss although the effects of natural ventilation can either assist or retard it depending on the time of year.

 

When air flows over a surface, the air molecules immediately next to the surface are at a standstill and those adjacent slip over those at rest with a resistance which is dependent on the viscosity of the air. A velocity gradient is formed where the velocity increases with increasing distance from the surface. The boundary layer created as a result of this phenomenon and the laminar sub-layer also formed as the boundary layer develops have a profound effect on the energy required to promote flow. Generally, the roughness of the surface of mine airways is large enough for the “bumps” to extend through the boundary sub-layer. The airway is then hydraulically rough and the resistance is a function of the relative roughness, i.e., the ratio of the roughness height to the diameter of the airway.

 

Most airways mined by conventional drill and blast techniques have roughness heights between 100 and 200 mm and even in very “blocky” ground, the average roughness height would not exceed 300 mm. Where airways are driven using boring machines, the roughness height is between 5 and 10 mm and still considered to be hydraulically rough. The roughness of airways can be reduced by lining them, although the justification is more usually ground support rather than a reduction in power required to circulate the ventilation air. For example, a large concrete-lined shaft with a roughness of 1 mm would be transitionally rough and the Reynolds number, which is the ratio of inertial to viscous forces, would also affect the resistance to airflow.

 

In practice, the difficulties in smooth concrete lining such a large shaft from the top down as it is being sunk results in increased roughness and resistances about 50% higher than the smooth values.

 

With a limited number of intake and return airways between the workings and surface, a large proportion (70 to 90%) of the total mine pressure loss occurs in them. Airway pressure losses also depend on whether there are any discontinuities causing shock losses such as bends, contractions, expansions or any obstructions in the airway. The losses resulting from these discontinuities such as bends into and out of airways, when expressed in terms of the losses which would be produced in an equivalent length of straight airway, can be a significant proportion of the total and need to be assessed carefully, particularly when considering the main intakes and exhausts. The losses in discontinuities depend on the amount of boundary layer separation; this is minimized by avoiding sudden changes in area.

 

Resistance of airways with obstructions

The effect of an obstruction on pressure losses depends on its drag coefficient and the fill coefficient, which is the ratio of the blockage area of the object and the cross-sectional area of the airway. The losses caused by obstructions can be reduced by minimizing boundary-layer separation and the extent of any turbulent wake by streamlining the object. Drag coefficients are affected by their shape and arrangement in the shaft; comparative values would be: I beam, 2.7; square, 2.0; cylinder, 1.2; elongated hexagon, 0.6; and fully streamlined, 0.4.

 

Even with small fill coefficients and low drag coefficients, if the obstruction is repeated regularly, such as with the beams separating hoisting compartments in a shaft, the cumulative effect on pressure losses is significant. For example, the resistance of a shaft equipped with semi-streamlined elongated hexagon beams and a fill coefficient of 0.08 would be about four times that of the concrete lined shaft alone. Although the material costs of the more readily available rectangular hollow structural steel sections are more than I beams, the drag coefficients are about one-third and easily justify their application.

 

Main and booster fans

Both axial and centrifugal fans are used to provide air circulation in mine ventilation systems, with fan efficiencies of over 80% being achievable. The selection between axial flow or centrifugal for main mine fans depends on cost, size, pressure, robustness, efficiency and any performance variation. In mines where a fan failure may result in dangerous methane accumulations, additional fan capacity is installed to ensure continuity of ventilation. Where this is not so critical and with a twin fan installation, about two-thirds of the mine airflow will continue if one fan stops. Vertical axial flow fans installed over the airways have low costs but are limited to about 300 /s. For larger air quantities, multiple fans are required and they are connected to the exhaust with ducting and a bend.

 

To obtain the highest efficiencies at reasonable cost, axial flow fans are used for low pressure (less than 1.0 kPa) applications and centrifugal fans for high pressure (greater than 3.0 kPa) systems. Either selection is suitable for the intermediate pressures. Where robustness is required, such as with exhausts with air velocities above the critical range, and water droplets are carried up and out of the system, a centrifugal fan will provide a more reliable selection. The critical air velocity range is between 7.5 m/s and 12.5 m/s where the water droplets may stay in suspension depending on their size. Within this range, the amount of suspended water can build up and increase the system pressure until the fan stalls. This is the region where some of the air recirculates around the blades and fan operation becomes unstable. Although not desirable for any type of fan, the possibility of a centrifugal fan blade failure is significantly less than an axial blade failure in this region of flow fluctuation.

 

It is rare that a main fan is required to operate at the same duty point over the life of the mine, and effective methods of varying fan performance are desirable. Although variable speed results in the most efficient operation for both axial and centrifugal fans, the costs, particularly for large fans, is high. The performance of an axial flow fan can be varied by adjusting the blade angle and this can be carried out either when the fan is stopped or, at a significantly higher cost, when it is rotating. By imparting a swirl to the air entering a fan using variable inlet vanes, the performance of a centrifugal fan can be varied while it is running.

 

The efficiency of the centrifugal fan away from its design point falls off more rapidly than that of an axial flow fan and, if a high performance is required over a wide range of operating points and the pressures are suitable, the axial flow fan is selected.

 

Ventilation systems

The position of the main fan in the overall system is normally on surface at the exhaust airway. The main reasons for this are simplicity where the intake is often a hoisting shaft and the exhaust is a separate single purpose airway and minimization of the heat load by excluding fans from intake airways. Fans can be installed at hoisting shafts either in forcing or exhausting mode by providing a sealed headframe. However, where workers, materials or rock also enter or leave the shaft, there is a potential for air leakage.

 

Push-pull systems where both intake and exhaust fans are installed are used either to reduce the maximum pressure in the system by sharing or to provide a very small pressure difference between the workings and surface. This is pertinent in mines using caving methods where leakage through the caved area may be undesirable. With large pressure differences, although air leakage through a caved zone is normally small, it may introduce heat, radiation or oxidation problems into the working places.

 

Underground booster fans, because of space limitations, are almost always axial flow and they are used to boost flow in the deeper or more distant sections of a mine. Their main drawback is the possibility of recirculation between the booster fan exhaust and the intake airways. By only providing a boost to the smaller airflows where they are required, they can result in a lower main fan pressure for the full mine airflow and a consequent reduction in total fan power required.

 

Secondary Ventilation

 

Auxiliary systems

Secondary ventilation systems are required where through ventilation is not possible, such as in development headings. Four arrangements are possible, each having its own advantages and disadvantages.

 

The forcing system results in the coolest and freshest air reaching the face and allows cheaper flexible duct to be used. The high velocity of the air issuing from the end of the supply duct creates a jet which entrains additional air and helps sweep the face of contaminants and provide an acceptable face velocity. Its main drawback is that the rest of the heading is ventilated with air that is contaminated with the gases and dust produced by mining operations in the face. This is particularly a problem after blasting, where safe re-entry times are increased.

 

An exhausting system allows all the face contaminants to be removed and maintains the rest of the heading in intake air. The drawbacks are that heat flow from the surrounding rock and moisture evaporation will result in higher face delivery air temperatures; operations in the heading back from the face, such as rock removal using diesel-powered equipment, will contaminate the intake air; there is no air jet produced to sweep the face; and more costly duct which is capable of sustaining a negative pressure is required.

 

In an exhaust-overlap system the problem of clearing the face with an air jet is overcome by installing a smaller fan and duct (the overlap). In addition to the extra cost, a disadvantage is that the overlap needs to be advanced with the face.

 

In a reversing system, the forcing ventilation mode is used, except during blasting and the re-entry period after blasting, when the airflow is reversed. Its main application is in shaft sinking, where re-entry times for deep shafts can be prohibitive if a forcing only system was used. The air reversal can be obtained by either using dampers at the fan inlet and outlet or, by taking advantage of a feature of axial flow fans, where changing the direction of blade rotation results in a flow reversal with about 60% of the normal flow being delivered.

 

Fans and ducts

The fans used for secondary ventilation are almost exclusively axial flow. To achieve the high pressures necessary to cause the air to flow through long lengths of duct, multiple fans with either contra-rotating or co-rotating impeller arrangements may be used. Air leakage is the greatest problem in auxiliary fan and duct systems, particularly over long distances. Rigid ducts fabricated from galvanized steel or fibreglass, when installed with gaskets, have suitably low leakage and may be used to develop headings up to several kilometres in length.

 

Flexible ducts are considerably cheaper to purchase and easier to install; however, leakage at the couplings and the ease with which they are ripped by contact with mobile equipment results in much higher air losses. Practical development limits using flexible duct rarely exceed 1.0 km, although they can be extended by using longer duct lengths and ensuring ample clearances between the duct and mobile equipment.

 

Ventilation Controls

Both through ventilation and auxiliary fan and duct systems are used to provide ventilation air to locations where personnel may work. Ventilation controls are used to direct the air to the working place and to minimize the short circuiting or loss of air between intake and exhaust airways.

 

A bulkhead is used to stop air flowing through a connecting tunnel. The materials of construction will depend on the pressure difference and whether it will be subject to shock waves from blasting. Flexible curtains attached to the surrounding rock surfaces are suitable for low pressure applications such as separating the intake and return airways in a room-and-pillar panel mined with a continuous miner. Timber and concrete bulkheads are suitable for higher pressure applications and may incorporate a heavy rubber flap which can open to minimize any blast damage.

 

A ventilation door is needed where pedestrian or vehicular passage is required. The materials of construction, opening mechanism and degree of automation are influenced by the pressure difference and the frequency of opening and closing. For high pressure applications, two or even three doors may be installed to create air locks and reduce leakage and the loss of intake air. To assist in opening air lock doors, they usually contain a small sliding section which is opened first to allow equalization of the pressure on both sides of the door to be opened.

 

A regulator is used where the amount of air flowing through a tunnel is to be reduced rather than stopped completely and also where access is not required. The regulator is a variable orifice and by changing the area, the air quantity flowing through it can also be changed. A drop board is one of the simplest types where a concrete frame supports channels into which timber boards can be placed (dropped) and the open area varied. Other types, such as butterfly louvres, can be automated and remotely controlled. On the upper levels in some open stoping systems, infrequent access through the regulators may be required and horizontally stiffened, flexible panels can be simply raised or lowered to provide access while minimizing blast damage. Even piles of broken rock have been used to increase the resistance in sections of a level where there is temporarily no mining activity.

 

Refrigeration and Cooling Systems

The first mine refrigeration system was installed at Morro Velho, Brazil, in 1919. Since that date, the growth in worldwide capacity has been linear at about 3 megawatts of refrigeration (MWR) per year until 1965, when the total capacity reached about 100 MWR. Since 1965 the growth in capacity has been exponential, with a doubling every six or seven years. The development of mine refrigeration has been influenced both by the air conditioning industry and the difficulties of dealing with a dynamic mining system in which the fouling of heat exchanger surfaces may have profound effects on the amount of cooling provided.

 

Initially, the refrigeration plants were installed on surface and the mine intake air was cooled. As the distance underground from the surface plant increased, the cooling effect was reduced and the refrigeration plants were moved underground closer to the workings.

 

Limitations in underground heat rejection capacity and the simplicity of surface plants has resulted in a move back to the surface location. However, in addition to the intake air being cooled, chilled water is now also supplied underground. This may be used in air-cooling devices adjacent to the working areas or as the service water used in drills and for dust suppression.

 

Refrigeration plant equipment

Vapour compression refrigeration systems are exclusively used for mines, and the central element of the surface plant is the compressor. Individual plant capacities may vary between 5 MWR and over 100 MWR and generally require multiple compressor systems which are either of the centrifugal or positive displacement screw design. Ammonia is normally the refrigerant selected for a surface plant and a suitable halocarbon is used underground.

 

The heat required to condense the refrigerant after compression is rejected to the atmosphere and, to minimize the power required to provide the mine cooling, this is kept as low as practical. The wet bulb temperature is always less than or equal to the dry bulb temperature and consequently wet-heat rejection systems are invariably selected. The refrigerant may be condensed in a shell and tube or plate and frame heat exchanger using water and the heat extracted and then rejected to the atmosphere in a cooling tower. Alternatively, the two processes can be combined by using an evaporative condenser where the refrigerant circulates in tubes over which air is drawn and water is sprayed. If the refrigeration plant is installed underground, mine exhaust air is used for heat rejection unless the condenser water is pumped to surface. Operation of the underground plant is limited by the amount of air available and higher underground wet bulb temperatures relative to those on surface.

 

After passing the condensed refrigerant through an expansion valve, the evaporation of the low temperature liquid and gas mixture is completed in another heat exchanger that cools and provides the chilled water. In turn, this is used both to cool the intake air and as cold service water supplied to the mine. The contact between water, ventilation air and the mine reduces water quality and increases heat exchanger fouling. This increases the resistance to heat flow. Where possible, this effect is minimized by selecting equipment having large water side surface areas that are easy to clean. On surface and underground, spray chambers and cooling towers are used to provide the more effective direct contact heat exchange between the air being cooled and the chilled water. Cooling coils which separate the air and water streams become clogged with dust and diesel particulate and their effectiveness rapidly declines.

 

Energy recovery systems can be used to offset the costs of pumping the water back out of the mine and pelton wheels are well suited to this application. The use of cold water as service water has helped to ensure that cooling is available wherever there is mining activity; its use has significantly improved the effectiveness of mine cooling systems.

 

Ice systems and spot coolers

The cooling capacity of 1.0 l/s of chilled water supplied underground is 100 to 120 kWR. On mines where large amounts of refrigeration are required underground at depths greater than 2,500 m, the costs of circulating the chilled water can justify replacing it with ice. When the latent heat of fusion of the ice is taken into account, the cooling capacity of each 1.0 l/s is increased approximately fourfold, thus reducing the mass of water that needs to be pumped from the mine back to surface. The reduction in pump power resulting from the use of ice to transport the coolness offsets the increased refrigeration plant power required to produce the ice and the impracticability of energy recovery.

 

Development is usually the mining activity with the highest heat loads relative to the amount of air available for ventilation. This often results in worksite temperatures significantly higher than those found with other mining activities in the same mine. Where the application of refrigeration is a borderline issue for a mine, spot coolers specifically targeted at development ventilation can defer its general application. A spot cooler is essentially a miniature underground refrigeration plant where the heat is rejected into the return air from the development and typically provides 250 to 500 kWR of cooling.

 

Monitoring and Emergencies

Ventilation surveys which include airflow, contaminant and temperature measurements are undertaken on a routine basis to meet both statutory requirements and to provide a continuing measure of the effectiveness of the ventilation control methods used. Where practical, important parameters such as main fan operation are monitored continuously. Some degree of automatic control is possible where a critical contaminant is monitored continuously and, if a pre-set limit is exceeded, corrective action can be prompted.

 

More detailed surveys of barometric pressure and temperatures are undertaken less frequently and are used to confirm airway resistances and to assist in planning extensions of existing operations. This information can be used to adjust the network simulation resistances and reflect the actual airflow distribution. Refrigeration systems can also be modelled and flow and temperature measurements analysed to determine actual equipment performance and to monitor any changes.

 

The emergencies that may affect or be affected by the ventilation system are mine fires, sudden gas outbursts and power failures. Fires and outbursts are dealt with elsewhere in this chapter and power failures are only a problem in deep mines where the air temperatures may increase to dangerous levels. It is common to provide a diesel-powered backup fan to ensure a small airflow through the mine under these conditions. Generally, when an emergency such as a fire occurs underground, it is better not to interfere with the ventilation while personnel who are familiar with the normal flow patterns are still underground.