Mine Planning Project

AHG University of science and Technology

Faculty of Mining and Geoengineering

Department of Economics and Management in Industry

Planned by

XYZ

Individual input data no. : 22

No.	Name	Unit	Value
1	Length of coalfield along strike, Ls	[m]	4000
2	Horizon depth, H	[m]	700
3	Horizon interval, h	[m]	150
4	Single longwall saleable daily production, Sld	[Mg/d]	7000
5	Overall productivity, P0	[Mg/emp/a]	800

Contents:

1. Introduction

2. Mine model layout

3. Reserves estimation

4. Mine construction schedule

5. Mine production schedule

6. CAPEX estimation

7. OPEX estimation

8. Evaluation of mine planning project efficiency

9. Sensitivity analysis

10. Summary – basic technical and economical parameters

Ad I

Introductory information and data:

Enterprise:

Method of mining:

Dip angle of deposit, α [degrees]:

Number of seams:

Thickness of seam I:

Thickness of seam II:

Thickness of seam III:

Distance between seam floors:

Methane emission:

Water inflow:

Number of longwalls, m:

Number of working days, N_s [d/a]:

underground hard coal mine

longwall caving

5

4 (enumerated: I, II, III)

3 [m]

3 [m]

4 [m]

20 [m]

11 [m³/Mg]

3 [m³/min]

2

250

Annual saleable production from longwalls, S_e [MMg/a]:

$$S_{e} = 10^{- 6}nS_{\text{ld}}N_{d} = \ 10^{- 6}*2*7000*250 = \mathbf{3,5}\ \lbrack\frac{\text{MMg}}{a}\rbrack$$

Annual saleable production from development workings, S_d [MMg/a]:

$$S_{d} = 0,1S_{e} = 0,1*3,5 = \mathbf{0,35\ }\lbrack\frac{\text{MMg}}{a}\rbrack$$

Annual saleable production of the mine, S [MMg/a]:

$$S = \ S_{e}*S_{d} = 3,5 + 0,35 = \mathbf{3,85}\ \lbrack\frac{\text{MMg}}{a}\rbrack$$

Overall manpower, N [man]:

$$N = \ \frac{10^{6}S}{P_{0}} = \ \frac{10^{6}*3,85}{800} = 4812,5\ \sim\mathbf{4813}\ \left\lbrack \text{man} \right\rbrack$$

Underground manpower, N_u [man]:

N_u = 0, 8N = 0, 8 * 4813 = 3850 [man]

Surface manpower, N_s [man]:

N_s = 0, 2N = 0, 2 * 4813 = 963 [ man]

u = 0, 8

Daily run-of-mine production, W_d [ThMg/d]:

$$W_{d} = \ \frac{1000S}{uN_{d}} = \frac{1000*3,85}{0,8*250} = \mathbf{19,25}\ \left\lbrack \frac{\text{ThMg}}{d} \right\rbrack$$

Underground productivity, P_u [Mg/man/d]:

$$P_{u} = \frac{1000W_{d}}{N_{u}} = \ \frac{1000*19,25}{3850} = \mathbf{5}\ \lbrack\frac{\frac{\text{Mg}}{\text{man}}}{d}\rbrack$$

Ad II

Mine model layout

1. Main shaft

2. Pit bottom of main shaft

3. Main transportation crosscut

4. Main transportation drift

5. Inclined drift

6. Main ventilation drift

7. Ventilation blind pit

8. Main ventilation crosscut

9. Pit bottom of air shaft

10. Air shaft

Main shaft depth, D_ms [m]:

D_ms = H + 20 = 700 + 20 = 720 [m]

Ait shaft depth, D_as [m]:

D_as = H − h + 10 = 700 − 150 + 10 = 560 [m]

Ad III

Reserves estimation

Seam number	Area [m^2]	Thickness [m]	Volume [Mm^3]	Density [Mg/m^3]	Proved reserves [MMg]	Coefficient ηav	Saleable reserves [MMg]
I	6 884 240	3	20,653	1,3	26,849	0,6	16,109
II	5 966 320	3	17,899	1,3	23,269	0,6	13,961
III	5 048 440	4	20,194	1,3	26,252	0,6	15,751
Total	17899000	10	58,74544		76,369		45,821

Area – calculated for every seam by multiplication of coal field length along strike by coal strike seam length along dip.

As you can see to calculate area for every seam we need to find x – coal strike seam length along dip.

$$\sin 5 = \ \frac{h_{\text{si}}}{x_{i}}\ \rightarrow x_{i} = \ \frac{h_{\text{si}}}{\sin 5}$$

For seam I:

h_s1 = 150[m]

$$x_{1} = \frac{h_{s1}}{\sin 5} = \ \frac{150}{\sin 5} \approx 1\ 712,06\lbrack m\rbrack$$

area for seam number I = H *  x₁ = 4000 *  1 721, 06 = 6 884 240[m²] 

For seam II:

h_s2 = 150 − 20 = 130 [m]

$$x_{2} = \frac{h_{s2}}{\sin 5} = \ \frac{130}{\sin 5} \approx 1\ 491,58\lbrack m\rbrack$$

area for seam number II = H *  x₂ = 4000 * 1 491, 58  = 5 966 320 [m²] 

For seam III:

h_s3 = 130 − 20 = 110 [m]

$$x_{3} = \frac{h_{s3}}{\sin 5} = \ \frac{110}{\sin 5} \approx 1\ 262,11\lbrack m\rbrack$$

area for seam number III = H *  x₃ = 4000 * 1 262, 11  = 5 048 440 [m²] 

Average coefficient of reserves recovery , η_av; let η_av = 0,6

Index of useful capacity of deposit, Z_u [Mg/m²]:

$$Z_{u} = \ \frac{1,3*\sum_{}^{}m_{i}}{\cos \propto}\eta_{\text{av}} = \ \frac{1,3*(3 + 3 + 4\ )}{\cos 5}*0,6 = \mathbf{7,83\ }\lbrack\frac{\text{Mg}}{m^{2}}\rbrack\ $$

Where:

m_i – I-th seam thickness [m],

α – dip angle [degrees]

Ad IV

No	Name	Unit	Depth/Length	Driving through	Rate of advance [m/mth]	Time [mth]
1	Preparatory workings before main shaft sinking					6
2	Main shaft sinking (1)	[m]	720	stone	60	12
3	Main shaft pit bottom inlet			stone		3
4	Driving the main transportation crosscut (3)	[m]	50	stone	50	1
5	Crossing (3/4L)	[m]		stone/coal		0,25
6	Driving the main transportation drift (4L)	[m]	1000	coal	150	6,7
7	Crossing (4L/5L)	[m]		coal		0,25
8	Driving the inclined drift I (5L)	[m]	1720	coal	120	14,3
9	Crossing (5L/6L)	[m]		coal		0,25
10	Preparatory workings before air shaft sinking					4
11	Air shaft sinking (10)	[m]	560	stone	60	9,3
12	Air shaft pit bottom inlet			stone		2
13	Driving the main ventilation crosscut (8)	[m]	50	stone	50	1
14	Crossing (8/6L)	[m]		stone/coal		0,25
15	Driving the main ventilation drift (6L)	[m]	1000	coal	150	6,7
16	Driving the main panel entry for longwall I	[m]	1000	coal	200	5
17	Driving the tail panel entry for longwall I	[m]	1000	coal	200	5
18	Driving the set-up entry for longwall I	[m]	300	coal	100	3
19	Longwall I face equipment installation					3
20	Start up of longwall I extraction
21	Driving the main transportation drift (4R)	[m]	1000	coal	150	6,7
22	Crossing (4R/5R)	[m]		coal		0,25
23	Driving the inclined drift II (5R)	[m]	1720	coal	120	14,3
24	Driving the main ventilation drift (6R)	[m]	1000	coal	150	6,7
25	Crossing (5R/6R)	[m]		coal		0,25
26	Driving the main panel entry for longwall II	[m]	1000	coal	200	5
27	Driving the tail panel entry for longwall II	[m]	1000	coal	200	5
28	Driving the set-up entry for longwall II	[m]	300	coal	100	3
29	Longwall II face equipment installation					3
30	Start up of longwall II extraction

Mine construction schedule

Ad V

Mine production schedule

Seam number	Saleable reserves [MMg]	Annual saleable production [MMg/year]	Production period [years]
I	16,109	3,85	4,2
II	13,961	3,85	3,6
III	15,751	3,85	4,1

Years:

Ad VI

1. Underground part

1. Expenditure on shaft sinking, I_u1 [MPLN]

No	Specification	Drive through	Size	Unit cost [ThPLN/unit]	Total [MPLN]
			unit	amount
1	Main shaft sinking (1)	stone	[m]	720	340
2	Driving the main shaft pit bottom inlet	stone		1
3	Air shaft sinking (10)	stone	[m]	560	250
4	Driving air shaft pit bottom inlet	stone		1
	Total	386,8

1. Expenditure on driving main development openings, I_u2 [MPLN]

No	Specification	Drive through	Size	Unit costs [ThPLN/unit]	Total [MPLN]
			Unit	Amount
1	Driving the main cross transportation opening (3)	stone	[m]	50	25
2	Crossing (3/4L)	stone/coal	[m]
3	Driving the main transportation drift (4L)	coal	[m]	1000	15
4	Crossing (4L/5L)	coal	[m]
5	Driving the inclined drift I (5L)	coal	[m]	1720	12
6	Crossing (5L/6L)	coal	[m]
7	Driving the main cross ventilation opening (8)	stone	[m]	50	25
8	Crossing (8/6L)	stone/coal	[m]
9	Driving the main ventilation drift (6L)	coal	[m]	1000	15
10	Driving the main longwall I panel entry	coal	[m]	1000	10
11	Driving the tail longwall I panel entry	coal	[m]	1000	10
12	Driving the set-up entry for longwall I	coal	[m]	300	15
13	Driving the main transportation drift (4R)	coal	[m]	1000	15
14	Crossing (4R/5R)	coal	[m]
15	Driving the inclined drift II (5R)	coal	[m]	1720	12
16	Driving the main ventilation drift (6R)	coal	[m]	1000	15
17	Crossing (5R/6R)	coal	[m]
18	Driving the main longwall II panel entry	coal	[m]	1000	10
19	Driving the tail longwall II panel entry	coal	[m]	1000	10
20	Driving the set-up entry for longwall II	coal	[m]	300	15
	Total	153,08

1. Expenditure on driving pump room and other chambers at pit bottoms, I_u3 [MPLN]

$$I_{u3} = aW_{d} + b + c\sqrt{Q_{w}} + dQ_{w} = 4,73*19,25 + 7,7 + 2,1\sqrt{3} + 2,7*3 = \ \mathbf{110,49}\ \lbrack MPLN\rbrack$$

where:

W_d – daily run-of-mine productions [ThMg/d]

Q_w – natural water inflowe in the mine [m³/min]

For the gassy mine without water problems: a=4,73; b=7,7; c=2,1; d=2,7

1. Expenditure on underground equipment (in gassy mine), I_u4 [MPLN]

$$I_{u4} = \left( \frac{40,5}{Z_{u}} + 16,2 \right)W_{d} = \left( \frac{40,5}{7,83} + 16,2 \right)*19,25 = \mathbf{411,42\ }\ \lbrack MPLN\rbrack$$

where:

Z_u – index od useful capacity of deposit [Mg/m²]

1. Expenditure on mechanization and automation (for gassy mine), I_u5 [MPLN]

I_u5 = 6, 9W_d + 31, 5 = 6, 9 * 19, 25 + 31, 5 = 164, 33 [MPLN]

where:

W_d – daily run-of-mine productions [ThMg/d]

Total expenditure on underground workings:

$$I_{u} = \ \sum_{i = 1}^{i = 5}{I_{\text{ui}} = 386,8 + 153,08 + 110,49 + 411,12 + 164,33 = \mathbf{1226,12}\ \left\lbrack \text{MPLN} \right\rbrack\ }$$

1. Surface part:

1. Expenditures on surface installations (e.g. winding installations, shaft head gear settings, construction of winding machine, processing plant, housing and other pit top buildings, etc.), I_s1 [MPLN]

I_s1 = 54 [MPLN]

1. Expenditure for surface coal processing and handling installations, I_s2 [MPLN]

I_s2 = 0, 45W_d(V_jbK_jb+C_mK_jm+M_mK_js) = 0, 45 * 19, 25(140*0,15+3,4*8,4+6*2,4) =  554, 05 [MPLN]

where:

W_d – daily run-of-mine productions [ThMg/d]

V_jb – processing/coal handling capacity of plant [m³/Mg/h]

K_jb – unit cost per m³ plant capacity [PLN/m³/Mg/h]

C_m – capacity od the coal handling machinery [Mg/Mg/h]

K_jm – unit costs of installation of processing machinery [PLN/(Mg/Mg/h)]

M_m – installed power of the processing plant [kW/Mg/h]

K_js – unit costs of the installed power [MPLN/(kW/Mg/h)]

Let: V_jb=140; K_jb=0,15; C_m=3,4; K_jm=8,4; M_m=6; K_js=2,4

1. Expenditure for preparation of mine stockyard and its fencing, I_s3 [MPLN]

I_s3 = aL_b + br + cn_z + dF + e + 5, 25 = 0, 45 * 0, 45 + 0, 1 * 20 + 0, 4 * 11 + 0, 45 * 12 + 0, 45 + 5, 25 =  17, 7 [MPLN]

where:

L_b – length of railway truck in mine stockyard area [km]

r – number of railway crossing in mine stockyard area,

n_z – number of loading track for coal,

F – surface area of the mine stockyard [ha]

Let: L_b=0,45; r=20; a=0,45; b=0,1; c=0,4; n_z=11; d=0,45; e=0,45; F=12

1. Expenditure for installation of hydraulic backfilling, I_s4 [MPLN]

No need for backfilling

I_s4 = 0 

1. Expenditure of the main ventilation infrastructure, I_s5 [MPLN]

I_s5 = (b_p+0,0065q_m)W_d + 10, 2S_w + 1, 5S_x + 13, 5n_m = (0,99+0,0065*6,6) * 19, 25 + 10, 2 * 1 + 1, 5 * 1 + 13, 5 * 1 = 45, 08 [MPLN]

where:

b_p –coefficient, b_p=0,09*q_s → b_p=0,99

q_{s –} volume of methane emission per Mg of coal extracted [m³/Mg]

q_m – yield of gas from methane emission points [m³/Mg]

S_w – number of exhaust ventilation shafts

S_x – number of intake ventilation shafts

N_m – number of methane emission points = number of exhaust shafts

Let: q_s=11; q_m=6,6; S_w=1; S_x=1; N_m=1

1. Expenditure on the installation of electric energy supply infrastructure, I_s6 [MPLN]

I_s6 = 12, 2 [MPLN]

1. Expenditure on the compressed air supply system, I_s7 [MPLN]

I_s7 = 1, 47W_d + 2, 25n_s1 = 1, 47 * 19, 25 + 2, 25 * 1 = 30, 55 [MPLN]

where:

n_s1 – number of compressed air stations = number of winding shafts = 1

1. Expenditure for hot water supply system, I_s8 [MPLN]

$$I_{s8} = \left( \frac{a}{P_{u}} + b \right)W_{d} + c = \left( \frac{2,5}{5} + 0,7 \right)*19,25 + 8 = \ \mathbf{31,1}\ \lbrack MPLN\rbrack$$

where:

W_d – as above

Pu – underground productivity [Mg/emp/d]

a, b, c – coefficients depending upon the supply model; for the projected mine model:

a=2,5; b=0,7; c=8

1. Expenditure on the constructions of administrative buildings and installation of telecommunication network, I_s9 [MPLN]

$$I_{s9} = \left( \frac{1,83}{P_{u}} + 0,13 \right)W_{d} + 3,8 = \left( \frac{1,83}{5} + 0,13 \right)*19,25 + 3,8 = \mathbf{13,35}\ \lbrack MPLN\rbrack$$

where:

W_d – as above

1. Expenditure on the supply centers, workshops and storage yards and narrow gauge track layout for mine, I_s10 [MPLN]

I_s10 = 2, 4W_d + 2, 63 = 2, 4 * 19, 25 + 2, 63 = 48, 83 [MPLN]

where:

W_d – as above

1. Expenditure on the mine premises preparation, land work, area protection and fencing, I_s11 [MPLN]

I_s11 = 16, 5 [MPLN]

1. Cost budgeting for mine development construction, I_s12 [MPLN]

I_s12 = 0, 4I_u2 = 0, 4 * 153, 08 = 61, 23 [MPLN]

where:

I_u2 – expenditure on main development openings [MPLN]

1. Reserve fund for unexpected expenditure, I_s13 [MPLN]

$$I_{s13} = 0,1\left( I_{u} + \sum_{i = 1}^{i = 12}I_{\text{si}} \right) = 0,1\left( 1226,12 + 884,6 \right) = \ \mathbf{211,07}\ \lbrack MPLN\rbrack$$

where:

I_u – total expenditure for underground mine development [MPLN]

1. Total expenditure on surface workings, I_s [MPLN]

$$I_{s} = \ \sum_{i = 1}^{i - 13}I_{\text{si}} = \ \mathbf{1095,67}\ \lbrack MPLN\rbrack$$

1. Total CAPEX – total expenditure (underground and surface), I [MPLN]

I = I_u + I_s = 1226, 12 +  1095, 67  = 2321, 78 [MPLN]

1. Calculation of annual amortization, A_an [MPLN/a]

$$A_{\text{an}} = \ \frac{I}{t_{\text{av}}} = \frac{2321,78}{18} = \mathbf{128,99\ }\lbrack\frac{\text{MPLN}}{a}\rbrack$$

where:

t_av – average period of influence of the mine assets, t_av = 18 years

1. Calculation of unit amortization, a [PLN/Mg]

$$a = \frac{A_{\text{an}}}{S} = \frac{128,99}{3,85} = \mathbf{33,50}\ \lbrack\frac{\text{PLN}}{\text{Mg}}\rbrack$$

where:

S – annual saleable production of the mine [MPLN/a]

Ad VII

Annual operating cost specification

ld	Cost specification	Fixed [MPLN/a]	Variable [MPLN/a]	Total [MPLN/a]	Unit cost [PLN/Mg]
1	Amortization	128,99		128,99	33,50
2	Materials	25,70	82,78	108,48	28,18
3	Energy	37,40	27,72	65,12	16,91
4	Equipment lease and rental	10,80		10,80	2,81
5	Drilling and mining services	7,10	55,83	62,93	16,34
6	Methane drainage services	3,00	23,49	26,49	6,88
7	Mining damage services	12,00		12,00	3,12
8	Other mining services	5,00	10,40	15,40	4,00
9	Repair services	25,00	15,40	40,40	10,49
10	Transport services	3,80	13,48	17,28	4,49
11	Other services	10,00	3,47	13,47	3,50
12	Labor costs	577,50		577,50	150,00
13	Welfare securities	117,81		117,81	30,60
14	Union benefits	45,05		45,05	11,70
15	Real property tax	7,65		7,65	1,99
16	Royalties		0,08	0,08	0,02
17	Environment charge		0,04	0,04	0,01
18	PFRON charge	3,10		3,10	0,81
19	Other taxes and charges	0,42		0,42	0,11
20	Insurances	4,60		4,60	1,19
21	Total production cost	1024,91	232,67	1257,58	326,64
22	Cost of sales			251,52	65,33
23	Total operating cost			1509,10	391,97
24	Cos minus amortization			1380,11	358,47

where:

S – annual saleable production of the mine [MMg/a], S=3,85

To point 12. let average annual salary is equal to 0,12 [MPLN/a] , N-manpower = 4813

Ad VIII

Evaluation of mine planning project efficiency

Year	0	1	2	3	4	5	6	7	8	9	10	11	12	13	14	15	16	17	18	19
Discount coefficient, at	1	0,91	0,83	0,75	0,68	0,62	0,56	0,51	0,47	0,42	0,39	0,35	0,32	0,29	0,26	0,24	0,22	0,20	0,18	0,16
Annual sales, St [MMg/year]					0,77	1,93	2,31	3,85	3,85	3,85	3,85	3,85	3,85	3,85	3,85	3,85	3,85	3,85	3,85	3,85
Average coal price, Pt [PLN/Mg]	450	450	450	450	450	450	450	450	450	450	450	450	450	450	450	450	450	450	450	450
Revenues, Rt [MPLN/year]					346,5	866,25	1039,5	1732,5	1732,5	1732,5	1732,5	1732,5	1732,5	1732,5	1732,5	1732,5	1732,5	1732,5	1732,5	1732,5
Annual cost, Ct [MPLN/year]					1380,11	1380,11	1380,11	1380,11	1380,11	1380,11	1380,11	1380,11	1380,11	1380,11	1380,11	1380,11	1380,11	1380,11	1380,11	1380,11
Investments, It [MPLN/year]	110,56	221,12	331,68	442,24	663,37	331,68	221,12
Inventories, Bt [MPLN/year]					65		65													-130
Capital expenditures, Nt=It+Bt	110,56	221,12	331,68	442,24	728,37	331,68	286,12													-130
Rt-Ct					-1033,61	-513,86	-427,23	352,39	352,39	352,39	352,39	352,39	352,39	352,39	352,39	352,39	352,39	352,39	352,39	352,39
at(Rt-Ct)					-705,97	-319,07	-241,16	180,83	164,39	149,45	135,86	123,51	112,28	102,08	92,80	84,36	76,69	69,72	63,38	57,62
atNt	110,56	201,02	274,12	332,26	497,48	205,95	161,51													-21,26
at(Rt-Ct-Nt)	-110,56	-201,02	-274,12	-332,26	-1203,45	-525,01	-402,67	180,83	164,39	149,45	135,86	123,51	112,28	102,08	92,80	84,36	76,69	69,72	63,38	78,87

where:

at = (1 + r)^−t

r – discount rate, r=0,1

Ad IX

Criteria:

$$NPV = \sum_{t = 0}^{t = 19}{a_{t}\left( R_{t} - C_{t} - N_{t} \right)} = \ - 1614,88$$

I didn’t calculate IRR because NPV≠0

Sensitivity analysis for criteria NPV

%	Average coal price	NPV	%NPV
-10	405	-2440,79	-0,51
-5	427,5	-2027,83	-0,26
0	450	-1614,88	0,00
5	477,5	-1110,15	0,31
10	495	-788,96	0,51

%	CAPEX	NPV	%NPV
-10	2089,606075	-1444,69	0,11
-5	2205,695302	-1529,78	0,05
0	2321,784528	-1614,88	0,00
5	2437,873755	-1699,97	-0,05
10	2553,962981	-1785,06	-0,11

%	Annual operating cost	NPV	%NPV
-10	1242,097259	-803,64	0,50
-5	1311,102663	-1209,26	0,25
0	1380,11	-1614,88	0,00
5	1449,113469	-2020,49	-0,25
10	1518,118872	-2426,11	-0,50

Ad X

Summary:

1. Reserves:

• proved: 76,369 [MMg]

• saleable: 45,821 [MMg]

1. Number of seams: 3

2. Number of shafts: 2

3. Horizon depth: 700 [m]

4. Horizon interval: 150 [m]

5. Number of longwalls: 2

6. Annual saleable production: 3,85 [MMg/a]

7. Overall manpower: 4813 [man]

8. Underground manpower: 3850 [man]

9. Overall productivity: 800 [Mg/emp/a]

10. Underground productivity: 5 [ Mg/man/d]

11. Daily run-of-mine production: 19,25 [ThMg/d]

12. CAPEX: 2321,78 [MPLN]

13. OPEX – total: 1380,11[MPLN/a]

14. NPV, IRR: NPV= -1614,88; I didn’t calculate IRR because NPV≠0

15. Conclusions derived from sensitivity analysis: