### Article Index

#### Problem description

This article presents case of the small scale trial excavation utilizing adapted backfilling mining method. Focus herein is on mining method itself and excavation technology.

Underground mine is already developed to the depth of 100m where excavation will take place. Average copper grade is 1.3%, ore body morphology is lens-like or irregular. Rock mass is weak, limiting stable span of the stope to 3m and striped pillar of 4m is required between adjacent stopes. Purpose of the trial excavation is to expand knowledge of the underground conditions and determine possibilities for improvements of resource recovery.

#### Ore body    window.dojoRequire(["mojo/signup-forms/Loader"], function(L) { L.start({"baseUrl":"mc.us3.list-manage.com","uuid":"f9b87849feded52c032b00fe5","lid":"08c06ce15d","uniqueMethods":true}) })

In order to perform trial excavation, small portion of the total ore body is selected for method and technology testing. Initial plan is to provide ~8500t per month in duration of approximately one year of production in the trial excavation. Around 180.000t is total reserves in the trial excavation block.

Figure 1 Shape of total ore body

Figure 2 Detailed overview of ore body portion selected for trial excavation

#### Development

Mine is already operational and level L+1600m from where trial excavation is meant to start has been already developed. Some existing underground openings will be actively used during trial excavation in order to service and communication between stopes.

Figure 3 Main development openings for trial excavation

Underground mine is opened with inclined shaft that connects all levels that are active up to date. This will be main mine entrance for the trial excavation. Since it is equipped with winch hoisting system, these underground openings will be main hoisting openings.

Level L+1600m has connection with the inclined shaft and significant portion of underground openings already exist and some of those will be used as main development for the trial excavation. From the connection with inclined shaft to the trial stopes there will be main transportation route for ore haulage with draw point at the inclined shaft.

#### Underground mining method

Ore body shape is presented in Figure 4 and Figure 5 along with underground openings at the level L+1600m. Resources tonnage for the trial excavation can be satisfied by small portion of total resource that is illustrated in Figure 6.

Figure 4 Shape of total ore body; underground openings at the level L+1600m

Figure 5 Part of the ore body selected for the trial excavation

Figure 6 Part of the ore body selected for the trial excavation

Figure 7 and Figure 8 present the part of the ore body selected for trial excavation and the nearby underground drifts at the level L+1600m. Selected part of the ore body spreads up the level L+1610m at its top. It is divided into 4 parts with 2.5m height where trial excavation takes place.

Figure 7 Detailed overview of ore body portion selected for trial excavation

Figure 8 Detailed overview of ore body portion selected for trial excavation

Geotechnical conditions for underground mining are not favorable. Rock mass is weak and available geotechnical database is not reliable. Due to weak rock mass it is not possible to apply any open stoping mining method, and morphology of the ore body limits application of caving methods. Having all this in mind, it is necessary to apply mining method with backfilling. This mining method has to be adjusted to the conditions in deposit. Description of the method and stope formation is explained in following section.

First step towards creating basic condition for trial excavation is extension of the existing drifts at the level L+1600m to the location of the ventilation shaft and its excavation up to the surface (Figure 9).

Figure 9 Extension of the existing drift at the level L+1600m and ventilation shaft position

Next phase of development requires that directional drifts in hanging wall and footwall are excavated following the contour of the ore body (Figure 10). These drifts are mutually connected at their ends. By connecting the directional drifts main ventilation is created and next steps are excavation of backfill shafts and ore passes. Development of all levels is in same manner, directional drifts are following boundary between ore and waste rocks.

Figure 10 Directional drifts in hanging wall and footwall with necessary connection between them and existing openings at the level L+1600m

Figure 11 presents the formation of the stopes at the level L+1600m. Stopes are 2.5m and 3m wide. Between two adjacent stopes 4m wide strip pillar is left.

Figure 11 Stopes formation at the level L+1600m

Directional drifts footwall are excavated by following the ore-waste boundary, while hangingwall drifts are placed following the contour of the topmost level (due to the backfilling technology and safety). After reaching the end, place where waste rock is encountered, stope is formed by connecting the two directional drifts. Figure 12  and Figure 13 illustrate stope formation at the level L+1600m. Ore is blasted in drifting manner and hauled to the ore pass. Advancing direction is from footwall towards the hanging wall.

Figure 12 Stope formation at the level L+1600m

Figure 13 Haulage of ore at the level L+1600m

At the level L+1600m all stopes are excavated in previously explained manner. After complete excavation of the stopes at this level, hanging wall drift is excavated up to the next level (L+1602.5m). Since it is excavated in ore, same is loaded and hanging wall directional drift is backfilled for the height of 2.5m. Then, excavation of the next level begins (Figure 14 and Figure 15). Crushed rock is used as backfilling material.

Figure 14 Typical situation at the stope; drilling is done by standing on the rock backfill

Figure 15 Ore is loaded from the level of footwall drift

Figure 16 Backfilling from the level of hangingwall drift

In general, excavation takes place in 2 Sectors (Figure 17). First sector has two parts, first part between footwall and hangingwall directional drifts with 11 stopes at each level, and second part outwardly directed from hangingwall drift with different number of stopes per each level. Second sector has 7 stopes between directional drifts.

Stope 0 is stope always excavated first and has direct connection with ventilation shaft.

Figure 17 Stope sequencing and markup

Selected mining method assumes that all stopes and pillars are strictly vertical and therefore stope positions are spatially same at each level. Only difference between levels comes from the fact the shape of the ore body is changed and length of stopes is adjusted accordingly. This is mostly case in the Sector 1, while stopes and drifts in the Sector 2 do not change position and are strictly above each other. This is caused by the shape and dip of the assumed ore body.

Figure 18 Stopes at the level L+1600m

Figure 19 Stopes at the level L+1602.5m

Figure 20 Stopes at the level L+1605m

Figure 21 Stopes at the level L+1607.5m

### Drilling and blasting

All underground excavation works are done using drill and blast technique. For drilling handheld pneumatic drill is used. Here, main drilling and blasting parameters are presented.

 Explosive Density: 1.15 g/cm3 Velocity of detonation: 4200m/s Patronne diameter: 28mm Type: PETN Rock properties Tensile strength: 5MPa Poisson’s ratio: 0.20 Drilling parameters Hole dimeter: 32mm Hole length: 1.6m Drilling speed: 0.2m/min

Burden is determined using expression:

$$B=\frac{0,17\ast P_h\ast r_h}{k\ast\sigma_t}$$

Where:

B – burden (m),

Ph – borehole pressure (GPa),

σt – Tenisle strength of rock (MPa),

k – coefficient based on Poisson’s ratio,

$$k=\frac{(1-\mu)}{(1+\mu)(1-2\mu)}=1.11$$

For the explosive with density higher than 1g/cm3 blasthole pressure is calculated as follows:

$$P_d=\frac{\rho_e\ast D^2}{8}$$

Where:

$\rho_e-$explosive density (g/cm3),

$D-$velocity of detonation (km/s).

Therefore:

$$P_d=\frac{1.15\ast{4.2}^2}{8}=2.5GPa$$

In case when explosive patron is smaller size than the borehole, blasthole pressure is reduce by following expression:

$$P_h=P_d\ast\left(\frac{d_e}{d_b}\right)^3$$

Where:

de – diameter of explosive patron,

db – diameter of borehole.

Therefore:

$$P_h=2.5\ast0.67=1.68GPa$$

And finally, burden is determined:

$$B=\frac{0.17\ast1680\ast0.016}{1.11\ast5}=0.85m$$

Blasting pattern for the stopes and drifts at other levels consists of 15 borehole of 1.6m length. Drilling is done using pneumatic handheld drill. Main drilling parameters in this case are:

 Number of boreholes 15 Length of borehole 1.6m Drilling speed 0.2m/min Explosive patrons per hole/face 4 / 60 Amount of explosive per hole 850g Amount of explosives per blast 12.75kg Specific amount of explosive 1.06kg/m3 / 0.35kg/t Initiation 500ms electric detonators

Figure 22 Blasting and initiation pattern for stopes and drifts at levels L+1602.5m, L+1605m, L+1607.5

Drilling time is:

$${{T}_{drill}}=\frac{15\cdot 1.6m}{0.2m/\min }\approx 120\min$$

Charging time is:

$${{T}_{ch}}=15\cdot 2\min =30\min$$

Total drilling and charging time is 150 minutes.

Table 1 Gantt chart for drilling and blasting for stopes and drifts at levels L+1602.5m, L+1605m, L+1607.5

Loading and hauling of blasted ore is done using Cavo 310 or similar machine type with following specifications:

 Weight 3150kg Body capacity 1m3 Bucket capacity 0.13m3 Net. loading capacity 0.7m3/min Tгavel speed 1-1.4m/s Аiг consumption, aveгage 133l/s Аiг pгessure гequiгement 4-7bar Min. tunnel dimensions 2620x2420mm (w x h)

Volume of ore produced by one blast:

$$V=1.4\cdot w\cdot h\cdot {{l}_{bh}}\cdot 0.95=3\cdot 2.5\cdot 1.6\cdot 1.4\cdot 0.95\approx 16{{m}^{3}}$$

Where:

W – width of stope (3m)

h – height of stope (2.5m)

lbh – length of blasthole (1.6m)

Total number of cycles necessary to transfer this amount of ore to the ore pass is:

$${{N}_{c}}=\frac{16{{m}^{3}}}{{{V}_{cb}}\cdot 0.9}=\frac{16}{1\cdot 0.9}\approx 18 cycles$$

Where:

Vcb – Volume of Cavo 310 body (1m3)

Each cycle consists of loading, driving to the ore pass, dumping and returning drive. Duration of one cycle is equal to the sum of durations of each separate operation.

$${{T}_{l}}=\frac{{{V}_{cb}}\cdot 0.9}{{{V}_{b}}}\cdot {{t}_{l}}=\frac{1{{m}^{3}}\cdot 0.9}{0.13{{m}^{3}}}\cdot 15s\approx 100s$$

Where:

Vcb – Cavo 310 body volume

Vb – bucket volume

Figure 23 Longest transportation distance at the level L+1600m

Total driving time can last up to:

$${{T}_{D}}=\frac{{{L}_{T}}}{\upsilon }=\frac{2\cdot 140m}{1m/s}=280s$$

Where:

LT – maximum total driving distance (2 x 140m see Figure 23)

$\upsilon$ – driving velocity (1m/s)

For dumping, maneuvering and all unexpected pauses 40s is left within each cycle and therefore total duration of one loading and hauling cycle is:

$${{T}_{LHC}}=280s+100s+40s=420s=7\min$$

Total duration of loading and hauling of one blasted face takes up to:

$${{T}_{LH}}={{T}_{LHC}}\cdot {{N}_{c}}=7\min \cdot 18=126\to 130\min$$

### Backfilling

Backfilling of stopes is performed by utilization of Cavo 310 load and haul machine. Maximum backfilling transportation distance is illustrated in Figure 24 and is around 125m at the level L+1600m.

Figure 24 Longest backfill distance

Volume of material that has to be installed in the stope:

$$V=w\cdot h\cdot {{l}_{bh}}=3\cdot 2.5\cdot 1.6\approx 12{{m}^{3}}$$

Where:

W – width of stope (3m)

h – height of stope (2.5m)

lbh – length of blasthole (1.6m)

Total number of cycles necessary to transfer this amount of ore to the ore pass is:

$${{N}_{c}}=\frac{12{{m}^{3}}}{{{V}_{cb}}\cdot 0.9}=\frac{12}{1\cdot 0.9}\approx 14 cycles$$

Where:

Vcb – Volume of Cavo 310 body (1m3)

$${{T}_{l}}=\frac{{{V}_{cb}}\cdot 0.9}{{{V}_{b}}}\cdot {{t}_{l}}=\frac{1{{m}^{3}}\cdot 0.9}{0.13{{m}^{3}}}\cdot 15s\approx 100s$$

Where:

Vcb – Cavo 310 body volume

Vb – bucket volume

Total driving time can last up to:

$${{T}_{D}}=\frac{{{L}_{T}}}{\upsilon }=\frac{2\cdot 125m}{1m/s}=250s$$

Where:

LT – maximum total driving distance (2 x 125m see Figure 24)

$\upsilon$ – driving velocity (1m/s)

For dumping, maneuvering and all unexpected pauses 40s is left within each cycle and therefore total duration of one loading and hauling cycle is:

$${{T}_{BFC}}=250s+100s+40s=390s\approx 6.5\min \to 7\min$$

Total duration of backfilling of one stope takes up to:

$${{T}_{BF}}={{T}_{BFC}}\cdot {{N}_{c}}=7\min \cdot 14=98\to 100\min$$

Table 3 Gantt chart for backfilling

#### Production capacity

Based on the times for technological phases and initial plan to achieve ~8500t of ore per month, following plan is created.

In each Sector two stopes are blasted and loaded per each shift, 4 stopes are operational at once for the whole mine. One blasting produces around 36t of the ore, therefore shift capacity is:

4 x 36t per stope = 144t per shift

With 2 production shifts per one day, and 22 working days per month, monthly production is:

2 shifts x 144t x 22 days = 6 336t per month