Numerical study on effects of coal properties on spontaneous heating in

longwall gob areas

Liming Yuan *,

*	 Corresponding author. Tel.: +1 412 386 4961; fax: +1 412 386 6595.
 

E-mail address: Lcy6@cdc.gov (L. Yuan).
 

Alex C. Smith

Pittsburgh Research Laboratory, National Institute for Occupational Safety and Health, P.O. Box 18070, Cochrans Mill Road, Pittsburgh, PA 15236, United States

abstract

A computational fluid dynamics (CFD) study was conducted to model effects of coal properties on the

potential for spontaneous heating in longwall gob (mined-out) areas. A two longwall panel district using

a bleeder ventilation system was simulated. The permeability and porosity profiles for the longwall gob

were generated from a geotechnical model and were used as inputs for the three-dimensional CFD modeling. The spontaneous heating is modeled as the low-temperature oxidation of coal in the gob using

kinetic data obtained from previous laboratory-scale spontaneous combustion studies. Heat generated

from coal oxidation is dissipated by convection and conduction, while oxygen and oxidation products

are transported by convection and diffusion. Unsteady state simulations were conducted for three different US coals and simulation results were compared with some available test results. The effects of coal

surface area and heat of reaction on the spontaneous heating process were also examined.

1. Introduction

Spontaneous combustion continues to be a problem for US

underground coal mines, particularly in western mines where

the coal is generally of lower rank. From 1990 to 1999, approximately 17% of the 87 total reported fires for US underground coal

mines were caused by spontaneous combustion [1]. The risk of

an explosion ignited by a spontaneous combustion fire is also present in those mines with appreciable levels of accumulated methane. In fact, three of the mine fires from the reported period

resulted in subsequent methane explosions. The incidence of such

fires is expected to increase with the projected increased mining of

lower rank coals, and the resulting explosion hazard is expected to

increase with the projected increased mining of deeper mines

(more methane) and the growth in the dimensions of longwall

panels (ventilation strain).

Spontaneous combustion of coal occurs when sufficient oxygen

is available and when the heat that is produced by the low-temperature reaction of coal with oxygen is not adequately dissipated by

conduction or convection, resulting in a net temperature increase

in the coal mass. Under conditions that favor a high heating rate,

the coal attains thermal runaway and a fire ensues. Spontaneous

combustion has also long been a problem in the storage and transport of coal. Much research has been done in experimental study

and mathematical modeling of spontaneous combustion of coals

as reviewed by Carras and Young [2]. The spontaneous combustion

potential of coals can be evaluated qualitatively in a laboratory

using one of four commonly used methods: adiabatic calorimetry,

isothermal calorimetry, oxygen sorption, and temperature differential methods. Although laboratory results are valuable, their

extrapolation to the mining environment has not been completely

successful because of complicated scaling effects that cannot be

reproduced in small scale experiments. Because of problems in

scaling of radiative heat transfer, the small scale spontaneous combustion results will not be able to be scaled to large-scale when

coal temperature is high enough that the radiative heat transfer

cannot be neglected. For small scale tests in which radiative heat

transfer can be neglected (low coal temperature), the test results

have also not been validated. The spontaneous heating of coal in

mines often occurs in a gob area and is not easily detected. The

amount of coal that accumulates in these areas and the degree of

ventilation can combine to give favorable conditions for spontaneous combustion. In order to prevent the occurrence of spontaneous

combustion in a gob area, it is important to investigate the spontaneous heating of coals under realistic mine ventilation conditions

with realistic methane generation and coal chemistry. Because of

the very nature and large size of a longwall gob area, it is very difficult to conduct full-scale tests. Thus, numerical modeling is

needed to assess spontaneous combustion susceptibility and provide insights for spontaneous heating for different coal properties

and ventilation conditions. Such modeling can also be completed

in a short time and at comparatively low cost.

Some numerical modeling studies have been done to understand the mechanisms of spontaneous combustion [3–10]. These

studies were one or two-dimensional models that mainly focused
on small size coal stockpiles. Little modeling work has been done

using actual underground mining conditions. Saghafi et al. did

numerical modeling of spontaneous combustion in underground

coal mines with a U-ventilation system [11,12], but their work

was also limited to two dimensions. Balusu et al. conducted a

CFD study of gob gas flow mechanics to develop gas and spontaneous combustion control strategies for a highly gassy mine [13]. In

this study, three-dimensional CFD modeling of spontaneous heating of different coals in longwall gob areas was conducted using

experimental data obtained by Smith and Lazzara [14] and Smith

et al. [15].

2. Gob layout and ventilation system

Longwall mining is an exploitation method used in flat-lying,

relatively thin tabular coal seams in which a long face is established to extract the coal. In these operations, a mechanical shearer

progressively mines a large block of coal, called a panel, which is

outlined with development entries or gate roads. The mined coal

is then hauled by a conveyor along the length of the face and removed from the mine. The roof is supported by special hydraulic

supports and shields that are advanced as the mining progresses.

The mine roof is allowed to cave behind in a carefully controlled

manner. The caved area is called the gob. The entry on the solid

coal block side of the reserve is called the headgate. The headgate

is used for the transportation of coal, men, and supplies into and

out of the face. The entry on the other side of the panel is used

for the passage of return air and is called the tailgate. When the panel entries on both sides have been developed, they are connected

by the entries at the end of the panel, called the bleeder entries. A

passageway driven between the entry and its parallel course is

called the crosscut. The airflow in an entry can be controlled by a

regulator that is usually constructed as an opening of adjustable

size in a ventilation door. Two types of ventilation systems are

used for longwall mining: referred to as bleeder and bleederless

ventilation systems. These ventilation systems are designed to

keep methane accumulation away from mining activities. In a bleeder system, the gob behind the active face, as well as previous

mined-out panels, is kept open and is ventilated to carry away

the methane gas from the gob area. In the bleederless system, seals

are installed in every crosscut after the panel is mined-out, and the

gob is isolated without any ventilation.

A typical longwall district in an underground coal mine may

consist of multiple panels. These panels are typically ventilated

using bleeder fans to ventilate the active and mined-out panels.

In mines with a demonstrated history of spontaneous combustion,

bleederless systems may be permitted as a spontaneous combustion control method. In the US, currently only two mines utilize

the bleederless system. In this paper, spontaneous combustion

with the bleeder ventilation system was studied, while in a later

phase of this study, spontaneous combustion with the bleederless

ventilation system will be investigated. In this work, two panels

are simulated, one as a mine-out panel and the other one as an active panel, utilizing a bleeder system and a bleeder fan. The layout

of the two panels and the ventilation system is shown in Fig. 1.

Each simulated gob area is 2000 m long, 300 m wide, and 10 m

high starting from the bottom of the coal seam. The ventilation airways are 2 m high and 5 m wide. Gob A represents the completed

panel, while gob B represents the active one. The ventilation

scheme includes a three-entry bleeder system. This ventilation

scheme and the panel dimensions are typical of longwall mines

operating in the Northern Appalachian Coal Basin of the Pittsburgh

coal seam. For western coal mines, the longwall panel layout and

ventilation systems may be slightly different from those in Pittsburgh coal seam, but the mechanisms that affect spontaneous

heating are the same. The simulation results obtained in this work

still address the typical spontaneous heating problems in western

coal mines. In the model, it is assumed that the middle entry between gob A and B and an entry on gob A’s tailgate side are partially open. All crosscuts between the first and second entries on

the headgate side of gob B are open as shown in Fig. 1. The bleeder

entries at the back end of the gob are represented as one entry connecting to the bleeder fan for modeling purposes. Fresh air at the

intake inlet is separated into four parts. Most of the air enters

the face to carry away methane from the working face and returns

to the exhaust fan. Some fresh air enters the gob of the active panel

B in the headgate side through the second intake entry. Some fresh

air travels through the third intake entry to dilute the methane

coming from the back ends of panel B and panel A. Gobs B and A

are ventilated by the bleeder fan. Except at the face, the air flow

in two gobs is towards the bleeder fan. Some fresh air enters the

tailgate entry of the panel A to dilute methane coming from gob

A. Four regulators are located at the end of the second and third intake entries and two tailgate entries, respectively, for controlling

the bleeder ventilation. The detailed ventilation data are described

in Section 4.

3. Modeling of low-temperature coal oxidation

The chemical reaction between coal and oxygen at low-temperatures is complex. Generally, three types of processes are believed

to occur [2]. These are (i) physical adsorption; (ii) chemical adsorption which leads to the formation of coal–oxygen complexes and

oxygenated carbon species; and (iii) oxidation in which the coal

and oxygen react with the release of gaseous products, typically

carbon monoxide (CO), carbon dioxide (CO2), and water vapor

(H2O). Of the above processes, oxidation is by far the most

exothermic.

The moisture content of coal can play an important role in the

low-temperature coal oxidation. The interaction between water

vapor and coal can be exothermic or endothermic depending on

whether the water condenses or evaporates. Sondreal and Ellman

[3] reported that for dried lignite, the rate of temperature increase

due to the adsorption of water increased with the moisture content

up to a value of 20% water (by mass) and then decreased with further increasing moisture content. Smith and Lazzara [14] found

that, initially, the rate of temperature rise depends on the heatof-wetting. Later the heating curves pass through an inflection

point, in which neither the heat-of-wetting mechanism nor the

oxidation mechanism dominates. In the final phase, the oxidation

mechanism dominates. The effect of the moisture content of the

air on the self-heating process was also dependent on coal rank

and temperature. Smith and Glasser [16] concluded that adsorption of water vapor does not in itself compete with the low-temperature oxidation in terms of ‘heat generation’, but appears to

speed up the oxidation rate, and possibly plays a catalytic role.

The same conclusion was reached by Smith and Lazzara [14]. In

underground coal mines, the fresh air is often saturated with moisture before it enters the gob. Although low rank coals usually contain more moisture, these moistures are not able to escape from

the coal into the saturated air. Therefore, the effect of moisture

on the spontaneous heating of the coal in the gob can be neglected.

The chemical reaction between coal and oxygen is simplified as

Coal þ O2 ! CO2 þ 0:1CO þ heat ð1Þ

The detailed chemical structure of coal varies with the rank and

origin of coal. According to experimental data [15], 1 mole of coal

reacting with 1 mole of oxygen generates 1 mole carbon dioxide

and roughly 0.1 mole carbon monoxide plus heat at the early stage

of coal oxidation. These stoichiometric data were for Wyoming No.
80 coal. The coal oxidation with oxygen is a type of heterogeneous

reaction. The heat of reaction is calculated using heat of formation

for reactants (coal), and products (CO2 and CO). The dependence of

the rate of oxidation, r, on temperature and oxygen concentration

can be expressed in the Arrhenius form:

r 1⁄4 A1⁄2O2]

n expð-E=RTÞ ð2Þ

where the chemical reaction rate is defined as the rate of change in

the concentrations of the reactants and products with a unit of

kmol/(m3 s), A is the pre-exponential factor with a unit of (kmol/

m3

)

1-n s-1

, E is the apparent activation energy with a unit of kcal/

mol, R is gas constant with a unit of kcal/(mol K), n is the apparent

order of reaction, T is the absolute temperature in K and [O2] is the

oxygen concentration with a unit of kmol/m3

. The value of the

apparent order of the reaction, n, in low-temperature oxidation

studies of coal and other carbonaceous materials has been shown

to vary from~ 0.5 to 1.0 [2,14], and is about 0.61 for some US coals

[17]. Using this value, the reaction rate becomes

r 1⁄4 1⁄2O2]

0:61A expð-E=RTÞ ð3Þ

The heat generated from oxidation is dissipated by conduction

and convection while the oxygen and oxidation products are transported by convection and diffusion. The convection in the gob is

caused by ventilation that creates a large pressure differential between the intake inlet and the bleeder fan. The detailed pressure

values at intake inlet and bleeder fan are described as part of the

boundary conditions in Section 4. The early stage of spontaneous

heating is a slow process, and the gas and coal particles are assumed to be in thermal equilibrium [2,9]. Therefore, the energy

transport equation is written as

oT oT oT oT u v ox þ oy

þ w oz ðeqgCpg þ ð1 eÞqcCpcÞ t þ qgCpg o

 ! o2

T o2T o2

T 1⁄4 keff 2 þ 2 þ 2 þ rQ ð4 ox oy oz Þ

 

 

-

where e is the porosity, qg, Cpg are the density and specific heat for the

gas, qc, Cpc are the density and specific heat for coal, Q is the heat of

reaction of coal oxidation and keff is the effective thermal conductivity of coal matrix. This effective thermal conductivity is calculated

as

keff 1⁄4 ekg þ ð1 - eÞkc ð5Þ

where kg and kc are the thermal conductivity for gas and coal.

The value of apparent activation energy, E, of different coals can

very between 12 and 95 kJ/mol. The pre-exponential factor, A, depends more on coal rank and measurement method. In this study,

the activation energy and pre-exponential factor data obtained by

Smith and Lazzara using an adiabatic heating oven were used [14].

regulator 4 bleeder fan A Tailgate 

intake B Headgate First intake Second Third intake entry intake entry entry regulator 1 

return 

Gob A (mined out) regulator 3 

Gob B (active) 

regulator 2 

face 

crosscut 

A Headgate 

B Tailgate 

300 m 

Fig. 1. Layout of gob areas and ventilation system.

In their study, minimum self-heating temperatures for 24 US coal

samples were determined and the activation energy and pre-exponential factor were derived using the simple Arrhenius equation

dT 1⁄4 A' expð-E=RTÞ ð6 dt Þ

This is a simplified form of Eq. (3). In this equation, the reaction

rate, r, is represented by the rate of temperature increase, and the

zero order reaction is assumed. The activation energy of coal is

independent of the order of reaction, while the pre-exponential

factor includes the dependence of reaction rate on oxygen concentration. Therefore, the activation energy in Eq. (6) is the same as in

Eq. (3), but the pre-exponential factor in Eq. (6) is different from

that in Eq. (3). In order to differentiate with the pre-exponential

factor A in Eqs. (3) and (6), here A* was used to denote the preexponential factor for the zero order reaction, thus it has a unit

of K/s. The relationship between A and A* can be obtained by an

applying the energy balance equation to coal particles

dT

qsC 0:61 ps d 1⁄4 Q1⁄2O2] A expð-E=RTÞ ð7

t Þ

where qs is the coal particle density in kg/m3 and Cps is the coal specific heat in J/(kg K). Substituting Eq. (6) into Eq. (7),

qsCpsA' expð-E=RTÞ 1⁄4 Q1⁄2O2]

0:61A expð-E=RTÞ

Thus

q C A 1⁄4 s ps A'

O 0 ð8Q1⁄2 2

Þ ] :61

In order to simulate the spontaneous heating of coal in longwall

gob areas, the source and amount of coal needs to be quantified.

Coal can be left from the mined coal seam or from other overlying

or underlying coal seams. In this study, the coal available in the

two gobs is from a 1-m thick rider coal seam that is originally

above the Pittsburgh coal seam. After the 2-m thick Pittsburgh coal

seam was completely mined-out (no coal left in the gob from this

coal seam), the rider coal seam caved into the bottom of the gob.

Fig. 2 shows the cross-section of two longwall panels with the caving coal layer at the bottom of the gobs. The spontaneous combustion can occur in this broken coal layer if oxygen is present. In gob

B, the air needed for the coal oxidation is from the leakage of ventilation air in the face entry and the second intake entry (the first

intake entry collapses and becomes part of the gob). In gob A,

the air supporting the oxidation is from the panel A tailgate entry

and the leakage from the gob B. Immediately above the broken coal

layer is the broken rock that forms the gob and is characterized by

gob porosity and permeability distributions. Because the broken

rock affects the gob flow, and the heat from coal heating can also

transfer to rock through convention and conduction, the entire
gob including broken rock and broken coal layer will be modeled. It

is assumed that the boundary of the gobs is adiabatic. In addition

to the two gobs, all the entries and crosscuts shown in Fig. 1 will

also be modeled.

The oxidation of coal will occur on any available coal surface,

including both external and internal pore surfaces, with oxygen

present. The reaction rate of spontaneous heating of coal in coal

stockpiles was found to be related to the external surface area

for nonporous coal particles with small pore diameters, and weakly

related or not related to particle size for small porous coal particles

with larger pore diameters [18]. Nugroho et al. [19] showed that

particle size has considerable influence on the self-heating character of coal. While a smaller particle reduces the critical ambient

temperature for spontaneous ignition to occur, the product of the

exothermicity and the pre-exponential factor QA, and the activation energy of the coals increase with decreasing particle size. Nugroho also reported that change of the critical ambient temperature

with particle size is almost negligible for porous coals, but significant for harder, nonporous coals.

It is difficult to define a coal particle size distribution in the coal

layer in the gob area because of the large gob size and different

caving characteristics in different locations of the gob. The critical

parameter that affects the heat generation and dissipation during

the spontaneous heating process in the modeling is the coal surface

area available in a unit volume, or surface-to-volume ratio. This

surface-to-volume ratio value roughly represents an average coal

particle size. The surface-to-volume ratio, s/v, for a spherical particle is 6/D with D the diameter of the particle, while the surface-tovolume ratio for a unit volume, S/V, is (1 - e)s/v with e the porosity

of the unit volume. So S/V = 6(1 - e)/D. If an average coal particle

diameter of 10 cm is assumed, then the surface-to-volume ratio

would be 36/m using a typical porosity value of 0.4 within the

uncompacted gob.

panel A middle crosscut 

second broken 

broken 

rock Gob A Gob B 

tailgate coal entry intake 

entry entry 

Fig. 2. Cross-section of two longwall panels with caving coal layer (not in scale).

4. Numerical modeling

In order to simulate spontaneous heating of coal in gob areas,

the steady state flow field in gobs was first modeled assuming that

the gob consists of only broken rocks. The basic equations solved

for the gas flow modeling are listed in Table 1.

Table 1

Basic equations solved for gas flow modeling

Mass

Momentum x-direction

Momentum y-direction

Momentum z-direction

Species

ou þ ov þ ow

ox oy oz 1⁄4 0

qðu ou þ v ou þ w ouÞ1⁄4- op þ lðo2u þ o2 u þ o2u l e e ox oy oz ox e u ox2 oy2 oz2 Þ - k

eqðu þ þ w Þ1⁄4- ov ov l v ov e op 2o

ox oy oz oy þ elð v

ox2 þ 2

þ 2 o v o v

oy2 oz2 Þ - k v

eqðu ow þ v ow op 2w 2w 2 ow o o o w l ox oy þ w oz Þ1⁄4-e oz þ elð ox2 þ oy2 þ oz2 Þ - k w

qu 2 oYi oY 2 2 i oYi o Yi o Yi o Yi

ox þ qv oy þ qw D oz 1⁄4 q i ox2 þ qDi y2 þ qDi z2 þ Si o o

Here l is kinetic viscosity; k is the permeability of the gob, D is the gas diffusion

coefficient, S is the rate of species production or consumption.

Then, the break coal

layer was added to the bottom of the gob replacing the original

broken rock at the same location. When spontaneous heating occurs in the broken coal layer, the steady state flow field will be affected, and becomes unsteady. Because the rock layer above the

coal layer affect both the gas flow and heat transfer, the entire

gob including both rock and coal layers will be simulated. The air

entering the gob form the intake inlet and panel A tailgate is fresh

air containing 21% of oxygen by volume. During spontaneous simulations, in addition to unsteady state mass, momentum and species (O2, CH4, CO, CO2) equations, energy Eq. (4) will also be solved.

A commercial CFD software program, FLUENT1

1 Reference to a specific product is for informational purposes and does not imply

endorsement by NIOSH.

from Fluent, Inc.,

was used in this study to simulate the gas flow and spontaneous

heating in the longwall gob areas. The gas flow in the longwall mine

gob area was treated as laminar flow in a porous medium using

Darcy’s law, while the gas flow in the ventilation airways was simulated as fully developed turbulent flow. The physical model and

mesh for the CFD simulation were generated using the mesh generator software, GAMBIT, from Fluent, Inc. The cell size varies from 1 to

4 m in the model with larger size cells located near the center of the

gob. The total cell number was over 1.5 million.

The permeability and porosity distributions of the gob areas

were generated using a geotechnical model: FLAC (Fast Lagrangian

Analysis of Continua), using the specific geological conditions of a

mine in the Pittsburgh coal seam [20]. In the FLAC modeling, mining was simulated in increments, starting from one side of the grid

and advancing to the other side. Extraction of the coalbed was

modeled by removing elements over the height of the coalbed.

The process of gob formation was modeled by first deleting rock

elements in the roof of the coalbed, so that they are stress relieved,

followed by inserting gob properties in these elements. Gob properties were also inserted in previously mined coalbed elements, so

that the gob filled the mined void. The gob properties such as elastic modulus, Poisson Ratio, cohesion and friction angle of different

rocks were measured in laboratory tests. The output of the FLAC

model is the permeability distribution in the gob. The model was

validated using empirical observation of the height of caving. The

permeability value varies from 2.97 x 10-8 to 8.42 x 10-7 m2

(3.0 x 104

–8.5 x 105 millidarcies), while the porosity value varies

from 0.17 to 0.41. Around the perimeter of the gob and immediately behind the face shields, the permeability and porosity values

were the largest, while near the center of the gob, these values

were the smallest. These permeability and porosity distributions

were then input into FLUENT.

Methane emission was also considered in the simulation because it affects the oxygen concentration distribution in the gob.

Methane is liberated from the face and inside the gob from the coal

remaining within and around the gob. Under normal mining conditions, most of the methane from the gob coal is released in the first

several hours, after which the release becomes slow. Since the time

period for spontaneous combustion is typically at least several

days, the gob coal methane emission is not considered in this simulation. Another source of continuous gob methane emission is

from any overlying or underlying coal seam reservoirs. The methane from these coal seam reservoirs is assumed to be released uniformly along the border between the gob and the reservoir. Using
ventilation data from a local Pittsburgh coal seam mine, the

amount of methane released from the rider coal seam reservoir

in this simulation is 0.13 m3

/s (280 cfm) for panel B and

0.024 m3

/s (50 cfm) for panel A. The methane emission rate from

the face to the panel B is 0.014 m3

/s (29 cfm). These data were also

entered into FLUENT using C subroutines.

The boundary conditions for ventilation pressures used in the

simulation were also obtained from a local Pittsburgh coal seam

mine’s ventilation data. The pressure was -0.7 kPa (-3.0 in. water

gauge) at the intake inlet, -0.87 kPa (-3.5 in. water gauge) at the

return outlet, and -2.7 kPa (-11.0 in. water gauge) at the bottom

of the bleeder shaft. The wall roughness of the ventilation airways

was adjusted to have a total intake airflow rate of 41 m3

/s

(87,000 cfm) in the active longwall panel. The pressure drops

through the two regulators located at the second and third intake

entries were also adjusted to have an airflow rate entering onto the

face of 28 m3

/s (60,000 cfm). The pressure drop at regulator 3 was

adjusted to have a flow rate at the return of 24 m3

/s (50,000 cfm),

and the flow rate in the entry on panel A’s tailgate side was 3.3 m3

/

s (7000 cfm) by adjusting the pressure drop at the regulator 4.

A simulation was conducted first without coal oxidation to obtain steady state flow field and gas distributions. Then, the unsteady simulation with coal oxidation was conducted using the steady

state solution as the initial conditions. The longwall face was assumed stationary during the simulation. The physical properties

for coal and air as input parameters are listed in Table 2.

Table 2

Physical properties for coal and air

Coal density 1300 kg/m3

Coal specific heat 1003.2 J/kg K

Coal conductivity 0.1998 W/m K

Heat of reaction 300 kJ/mol O2

Air specific heat 1.0 J/kg K

Air conductivity 2.6 x 10-2 W/m K

Air dynamic viscosity 1.8 x 10-5 kg/m s

Gas diffusion coefficient 1.5 x 10-5 m2

/s

Initial temperature 300 K

Fig. 3. Flow path lines colored by velocity magnitude (m/s) in gob areas.

Fig. 4. Contours of oxygen concentration (1 = 100%) in gob areas.

5. Flow patterns inside the gob

The flow patterns inside a gob will have a significant effect on

the spontaneous heating of coals because the oxygen needed for

the oxidation is provided by the gas flow, and the heat generated
from the oxidation may be carried away by the gas flow. The gas

flow inside a gob is expected to be three-dimensional with the flow

in the vertical direction, but weaker than in the other two directions due to reduced permeability and pressure gradients. In order

to visualize the flow patterns inside the gob, a virtual horizontal

reference surface was created 1 m from the bottom of the mined

seam floor to compare the results with respect to this horizontal

reference surface. Fig. 3 shows the flow path lines colored by velocity magnitude in the two gob areas. In the face entry, most air flows

from intake inlet to return outlet. Inside the gob, because of pressure differential between the intake inlet and the bleeder fan, air

flows from the headgate side to the tailgate side. At the face, air entered into the gob through the gaps between the face shields and

flowed to the middle entry between the two panels. The higher

gas velocity was between 1.5 x 10-4 and 3.0 x 10-4 m/s (0.03 to

0.06 fpm) near the back end of panel B. Fig. 4 shows the oxygen

distribution in two panels. The oxygen concentration was at 21%

in most of mined-out gob A. In the active gob, B, the oxygen concentration close to the intake entry was below 21% because of

methane emission from the overlying coal seam reservoir. The

openings of crosscuts between the first and second intake entries

appear to be very important to dilute the methane concentration

near the perimeter of the gob on the headgate side of the active panel. The purpose of opening these crosscuts is to bleed the gob

(sweep out the methane). As discussed later, these open crosscuts

at same time may facilitate the spontaneous combustion by providing enough oxygen. Therefore, in a bleederless ventilation system, all these crosscuts are completely sealed.

6. Simulation results and discussion

Three coals tested by Smith and Lazzara [14] were selected for

these simulations as the coal left in the gob areas. Three coals were

labeled as No. 80-1, E-1 and Pittsburgh coal, respectively. The adiabatic heating oven test results for these three coal samples are

shown in Table 3.

Table 3

Laboratory-scale test results for three coals [14]

Coal sample Minimum SHT, °C E, kcal/mol A, K/s

No. 80-1 35 15.9 1.9 x 106

E-1 65 17.6 1.1 x 107

Pittsburgh 90 21.1 4.4 x 108

Based on the laboratory-scale evaluation, coal

sample No. 80-1 has a low minimum self-heating temperature

(SHT) and activation energy representing a high tendency for spontaneous combustion. Pittsburgh coal sample has high minimum

SHT and activation energy, representing a lower potential for spontaneous combustion. The E-1 coal sample falls in between in terms

of spontaneous combustion potential. These values represent a relative spontaneous combustion potential. Under the right conditions, all coals can undergo spontaneous combustion.

Spontaneous heating can begin at ambient temperature when

coal is exposed to oxygen. As the self-heating proceeds, the coal

temperature increases slowly. The temperature rise usually consists of two stages. The first stage is a slow temperature rise while

the second one is a fast temperature rise. The start of the second

stage is also called thermal runaway. The time to reach a thermal

runaway is called induction time. When the coal temperature

reaches about 500 K (230 °C), the spontaneous heating mechanism

changes to rapid combustion [21]. So the simulations were focused

on the self-heating mechanism at temperatures below 500 K. The

surface-to-volume ratio used in the simulations is 36/m, which is

equivalent to an average coal particle diameter of 10 cm. Fig. 5

shows temperature and oxygen concentration distributions for

the No. 80-1 coal after 10 days. It is apparent that temperature

rises occurred in three areas. Area I is the area close to the active

tailgate and around the return. Area II is the area nearby the crosscuts and close to the back end of the active panel B, and area III is

close to the middle entry and back end of the mined-out panel A.

There was very little or no temperature rise in the other areas. This

is because oxygen was nearly all consumed by coal oxidation at the

periphery of the gob areas and no oxygen was available inside the

gobs. This is evidenced by the oxygen concentration distribution in

Fig. 5d and indicates that spontaneous heating is mainly controlled

by oxygen availability. Under the ventilation conditions studied

here, not enough oxygen could be carried to the compressed gob

area through convection and diffusion before being consumed on

the edges of the gob. The temperature rise rates in three areas were

also different. Inside each area, temperature was not uniform. At

any time, there existed a maximum temperature in each area.

From a fire safety point of view, the maximum temperature represents the worst scenario. Hereafter, the maximum temperature at

any time in each area was used to represent the temperature in

this area, respectively. Fig. 6 shows the temperatures versus time

in three areas. The temperature rises in areas I and III were about

2 days behind those in area II. Fig. 7 shows temperature and oxygen concentration distributions for the E-1 coal after 17.5 days.

Temperature rise also occurred only in these three areas mentioned previously. The main difference is that area II had a fast

temperature rise after about 14 days, while area I and III were still

in the slow increase stage after 18 days with the temperatures below 350 K as shown in Fig. 8. Oxygen was available close to the

face, but was depleted very quickly deeper into the gob. Around

area II, some oxygen, about 4%, was available, probably accounting

for the higher temperature there. Temperature and oxygen concentration distributions for Pittsburgh coal after 40 days are shown in

Fig. 9. Higher temperature occurred in areas II and III. Because the

temperature rise was very small, 4 degrees, oxygen was available

everywhere in the gob areas. This led to slight temperature rises

in other areas of the gob.

6.1. Maximum temperature

For all three coals, at any time, area II had the highest temperature. Hereafter, this temperature was referred as the maximum

temperature in the gob at given time. Fig. 10 shows the maximum

temperature versus time for the three coals using the surface-tovolume ratio of 36/m. For No. 80-1 coal, the maximum temperature rose slowly for the first 5 days. After that it rose very rapidly.

For E-1 coal, the maximum temperature rose rapidly after about

two weeks. For Pittsburgh coal, after 40 days, the maximum temperature rise was only several degrees. E-1 coal has a minimum

SHT that is between those for No. 80-1 and Pittsburgh seam coals.

In these simulations, it still reached a thermal runaway, but had a

10 day longer induction time compared with E-1 coal, while Pittsburgh coal does not reach a thermal runaway in 40 days. This indicates that for the same underground mining conditions, a higher

laboratory minimum SHT value means that the time to reach thermal runaway is longer.

It should be pointed out that the maximum temperature rise

was sensitive to the kinetic properties of the coals: activation energy and pre-exponential factor. The activation energy and preexponential factor data were obtained from the laboratory-scale

study using very fine coal particles with average diameter between

75 and 150 lm [14]. With larger coal particles, most research has

demonstrated that the pre-exponential factor will decrease, but

the exact relationship between coal particle size and pre-exponential factor is far from well understood. In a real situation with larger coal particles, the rate of temperature rise will be less than

those calculated here. However, the temperature data from this
study can represent results from a worst-case spontaneous combustion scenario.

Fig. 5. Temperature and oxygen concentration distributions for No. 80-1 coal after 10 days: (a) temperature in entire gob (K); (b) temperature in area I (K); (c) temperature in

areas II and III (K); (d) oxygen concentration (1 = 100%).

6.2. Comparison with large-scale test results

Smith et al. conducted large-scale studies of spontaneous heating of a large coal mass under conditions that simulate a gob area

of a mine [15]. An insulated coal chamber with dimensions of

4.5 m x 1.8 m x 1.8 m was used. The chamber held up to

11,794 kg of coal and was provided with a forced ventilation system. Three experiments were completed using a high-volatile C

bituminous coal that exhibited a high spontaneous combustion potential in laboratory-scale tests. In the first two tests, a sustained

heating was not achieved. In the third test, temperatures throughout the coalbed increased steadily from the start, with thermal

runaway occurring near the center of the coalbed after 23 days.

The thermal reaction zone then moved toward the front of the

coalbed due to oxygen depletion in the center of the coalbed. The

coal used in the third test was the No. 80-1 coal from Table 3. In

the experiment, the entire coalbed showed indications of heating

immediately after the airflow was started. The temperature histories of the coalbed were recorded using thermocouples arranged in

seven vertical arrays 0.3 m, 0.9 m, 1.5 m, 2.1 m, 2.7 m, 3.4 m, and

4.0 m from the front of the coalbed. Each array contained nine

thermocouples evenly distributed over the surface along the coalbed width direction. The maximum temperature appeared on one

of the thermocouples on the array 0.3 m from the front of the coalbed, the closest one, also because of maximum available oxygen.

Fig. 11 shows this maximum temperature history compared with

the maximum temperature in gobs obtained from simulations.

The temperature rise curves are very similar except that the induction times are different. In the test, it took about 25 days before the

thermal runaway occurred. The induction time discrepancy between modeling and test results may be attributed to two factors.

First, the dimensions of the large-scale test (4.5 m x 1.8 m x

1.8 m) are far smaller than those for the gob (2000 m x 300 m x

10 m), leading to much lower heat capacity for the coalbed chamber. The sidewalls and floor of the coalbed chamber were insulated

with firebricks, while the boundary of the gob was assumed completely adiabatic. Therefore, the relative heat loss from the coalbed

chamber boundary was much higher than heat loss from the

boundary of the gob (no heat loss), leading to longer time to the

thermal runaway. The second factor is the air flow velocity. The

average airflow velocity in the coalbed chamber was estimated

as about 3.3 x 10-3 m/s, while the airflow velocity inside the gob

was between 3 x 10-5 and 3 x 10-4 m/s. The higher airflow veloc­
ity carried away more heat generated from the spontaneous heating, also leading to longer time to reach the thermal runaway. In

general, the simulation results are consistent with the test results.

More field tests are needed to compare with the modeling results

in detail.

Fig. 5. (continued)

500 

450 

Area II Area I 

Area III 

Temperature (K) 400 

350 

300 

0 2 4 6 8 10

Time (day) 

Fig. 6. Temperature–time histories in three areas for No. 80-1 coal.

6.3. Effect of coal surface area

The effect of coal surface area on the spontaneous heating was

examined by using different surface-to-volume ratios, S/V, for No

80-1 coal. Fig. 12 shows the maximum temperature histories for

simulations using three surface-to-volume ratios: 36, 12 and 2 /

m. This is equivalent to an average coal particle diameter of 10,

30, and 180 cm, respectively. With S/V = 2, the average particle

diameter is 180 cm. In reality, this does not indicate a volume with

coal particles with a diameter of 180 cm. Instead, it represents a

volume with small coal particles coexisting with other non-coal

rock particles. When the S/V was reduced to 12/m, not only was

the induction time increased to about 17 days, but also the temperature rise in the second stage became less steep, and after about 5

days changed into another slow temperature rise stage, probably

because of less heat released in a unit volume. After the S/V was

further reduced to 2, maximum temperature rise was only a few

degrees in 35 days, and there was no thermal runaway. This indicates that decreasing the coal reaction surface area will greatly reduce the spontaneous combustion fire hazard.
Fig. 7. Temperature and oxygen concentration distributions for E-1 coal after 17.5 days: (a) temperature (K); (b) oxygen concentration (1 = 100%)

Temperature (K) 

500 

450 

400 

350 

300 

Area II 

Area III 

Area I 

0 5 10 15 20 

Time (day) 

Fig. 8. Temperature–time histories in three areas for E-1 coal.

6.4. Effect of heat of reaction

Fig. 13 shows simulated maximum temperature histories for

coals with different heats of reaction. Three heats of reaction were

chosen: 3.0 x 108

, 2.4 x 108 and 1.6 x 108 J/kmol O2; roughly

equivalent to the heating values of 13,125, 10,500, and 7000 Btu/

lb, respectively. The activation energy and pre-exponential values

for coal No. 80-1 were used. Coal with a heat of reaction of

3.0 x 108 J/kmol O2 had an induction time of 5 days, and coal with

a heat of reaction of 2.4 x 108 J/kmol O2 had an induction time of

about 7 days. Their maximum temperature–time curves are also

very similar. For the coal with a heat of reaction of 1.6 x 108 J/kmol O2, the induction time increased to about 12 days, and it took

about 20 days to reach 500 K, indicating that the heat of reaction

of coal oxidation does not have a major effect on spontaneous heating process.

7. Conclusions

CFD simulations were conducted to model the spontaneous

heating of three coals in longwall gob areas. These coals have different self-heating potentials based on their minimum SHTs evaluated in a laboratory-scale study. Simulation results demonstrate

that under typical longwall mining conditions, the spontaneous

heating hazards of different coals were strongly dependent on

mine ventilation conditions and coal properties. A coal with a larger minimum SHT would result in a longer induction time for

spontaneous combustion. Compared with the most active No. 80­

1 coal, the E-1 coal’s induction time increased by an additional
300

350

400

450

500

0 5 10 15 20 25

x108

 J/kmol

Time (day)

Maximum temperature (K)

3.0x108 J/kmol 2.4x108 J/kmol 1.6

10 days, while for Pittsburgh coal, there was no significant temperature rise in 40 days. The temperature rise occurred in only three

areas for No. 80-1 and E-1 coals. The maximum temperature in

gobs occurred in area II. For No. 80-1 coal, temperature rise in areas

I and III was about two-days behind that in area II, while for E-1

coal, temperature rise in areas I and III was still in the slow increase

stage when the temperature in area II was already above 500 K.

When the temperature rise was significant, the spontaneous heating became mainly oxygen controlled because oxygen originally

present in the gob was consumed by coal oxidation at the early

stage.

Fig. 9. Temperature and oxygen concentration distributions for Pittsburgh coal after 40 days: (a) temperature (K); (b) oxygen (1 = 100%).

500 

No.80-1 E-1 

Pittsburgh 

Maximum temperature (K) 450 

400 

300 

0 5 10 15 20 25 30 35 40 

Time (day) 

Fig. 10. Simulated maximum temperature–time histories in gobs for three coals.

350 

500 

simulation test result

Maximum temperature (K) 450 

400 

350 

300 

0 5 10 15 20 25 30 35 

Time (day) 

Fig. 11. Comparison of maximum temperature–time histories of No. 80-1 coal

between simulation and test results [15].

40 
300 

350 

400 

450 

500 

Maximum temperature (K) 

S/V: 36 S/V:12 

S/V: 2 

0 5 10 15 20 25 30 35 

Time (day) 

Fig. 12. Simulated maximum temperature-time histories for three surface-tovolume ratios: 36, 12 and 2 /m for No. 80-1 coal.

300 

350 

400 

450 

500 

Maximum temperature (K)

3.0x108 J/kmol 2.4x108 J/kmol 1.6x108 J/kmol 

0 5 10 15 20 25 

Time (day) 

Fig. 13. Simulated maximum temperature-time histories for No. 80-1 coal with

different heats of reaction.

The temperature rise was greatly dependent on the available

coal reaction surface area. For larger coal particle size, the reaction

surface area and the maximum temperature rise were both significantly reduced. The heat of reaction of coal oxidation had a minor

effect on spontaneous heating. With a lower value of heat of reaction, the induction time was increased but the fire hazard was still

present.

In real mining conditions, the face is moving at a certain speed.

This movement creates a dynamic flow field that greatly affects the

spontaneous heating in the gob area. More research is needed to

study the effect of the longwall face movement on the spontaneous

heating of coals in the gob areas. Future research also needs to take

into account of evaporation or condensation of coal moisture, and

the decrease of the rate of oxygen sorption with increasing uptake.

Modeling results from this study are consistent with available

large-scale test data except for different induction time. More

large-scale test data are needed in the future to further improve

the modeling of spontaneous heating in underground gob areas.

Acknowledgements

The authors wish to acknowledge the technical help from Drs.

Gabriel S. Esterhuizen and C. Özgen Karacan with the estimation

of permeability and porosity in the gob.

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