Optimizing Blasting in Narrow Veins: Using DECKs to Reduce Dilution and Improve Recovery in Underground Mining

Mining is one of the main economic activities worldwide and a fundamental pillar in the development of many countries [1, 2]. According to the World Bank, mining accounts for around 10% of global Gross Domestic Product (GDP) [3, 4] and is responsible for the provision of essential raw materials for various industries, such as energy, construction and technology [5-7]. In Latin America, mining is even more important, accounting for approximately 20% of the region's GDP, with countries such as Chile, Peru and Mexico standing out as major producers of copper, gold and silver [8-10]. In the specific case of Peru, mining is a key sector for the national economy, representing more than 10% of the GDP and being the main generator of exports, with minerals such as copper, gold and zinc as its main products [11-13]. The stability and sustainability of mining are essential not only for the economic development of the region, but also to ensure a constant supply of mineral resources globally [14].

Underground mining is one of the most complex forms of mineral extraction due to the geological and operational conditions it faces [15, 16]. Unlike open-pit mining, where ore layers are more directly accessible, underground operations requires specialized techniques to work in confined spaces and with high geotechnical pressure [17, 18]. In this type of mining, narrow veins, which are low-thickness mineralized formations, represent one of the greatest challenges [19]. These veins, by their nature, require precise mining methods to extract the mineral without affecting the stability of the rock mass that surrounds them [20-22]. Their complex geometry and high geomechanical variability make them difficult to access and recover efficiently, often resulting in a higher risk of ore dilution. Dilution occurs when, during the blasting or extraction process, waste material is mixed with the valuable ore, reducing the grade of the mined ore and negatively affecting the profitability of the operation [23, 24].

One of the main causes of ore dilution in underground mining is the use of excessive operating charges during blasting [25, 26]. This phenomenon can cause additional damage to the rock mass, compromising its stability and increasing the fracturing of the surrounding rock [27]. Excessive fracturing not only increases dilution but also results in more non-valuable material that must be processed, which increases operating costs and reduces the efficiency of the extraction process [26]. To mitigate this problem, several alternatives have been implemented in recent decades, such as reducing the number of explosives used, controlling sequencing of blasting, and optimizing drilling designs [25]. However, these approaches have not achieved significant results in terms of reducing dilution or improving rock mass stability. Explosive reduction, for example, has proven effective only in certain contexts, but does not always translate into a substantial decrease in fracturing or improved mineral recovery. Similarly, blast sequencing and adjustment of drilling designs sometimes fail to adequately control the effects of explosive energy on stope stability. 

In this context, the optimization of the operating charge in underground blasting has become a critical aspect for mining operations, especially those exploiting narrow veins. The proper selection and control of the operating charge not only minimizes rock mass fracturing but also reduces ore dilution and improves operational efficiency. This, in turn, translates into greater mineral resource recovery and a significant reduction in the costs associated with extraction and processing.

Therefore, this study aims to evaluate the use of DECKs as an alternative for the optimization of the operating charge in underground mining. The use of DECKs, which consists of inserting inert material between the explosive charges, has shown potential to better control the energy released during blasting, which in turn minimizes unwanted fracturing of the rock mass. Unlike other solutions that have been implemented in the past, such as reducing the number of explosives or sequencing blasts, the use of DECKs offers a more precise approach to controlling the explosive charge, allowing a more efficient use of the released energy. This method not only reduces dilution by avoiding over-setting of the rock but also improves the stability of the rock mass by reducing the unwanted effects of blasting on the structure of the pit. In addition, DECKs provide a significant benefit in terms of sustainability, as they use inert materials instead of additional explosives, which contributes to reducing the environmental impact associated with blasting.

Regarding the proposed solution, the use of DECKs in underground mining has proven to be an effective technique to optimize blasting processes and minimize ore dilution. By using DECKs, which consist of inserting inert material between the explosive charges, the energy released during detonation is better controlled, which allows reducing unwanted fracturing of the rock mass. This technique is particularly relevant in the exploitation of narrow veins, where precise control of the operating charge is crucial to avoid over-fracturing and the mixing of waste material with the valuable mineral. In addition, the use of DECKs improves the stability of rock mass, since the controlled distribution of explosive energy reduces adverse effects on the structure of the pit, which, in turn, increases operational safety and efficiency.

This study has a high potential in underground mining, especially in the Peruvian context, since, although there are several studies that address blast optimization and dilution reduction, none have focused specifically on the application of DECKs as an effective solution in narrow vein mining. Research of this type is particularly relevant, as it allows filling a gap in the literature on the practical application of this technique in specific geomechanical conditions, which could have a significant impact on the mining industry globally. Thus, Meng et al. [28] addressed the reduction of dilution in narrow vein mining, a critical challenge in underground mining. The research proposes a sequencing of operational strategies to improve drilling and blasting efficiency, comparing conventional methods such as cut and fill (CAF) with more productive methods such as longhole stoping. In this context, decoupled loading techniques were implemented in the blasting process, resulting in a significant reduction of dilution, from 40% to 8%, and a decrease in the Power Factor (PF) (kg/tn) by 25%. Furthermore, this approach allowed a reduction in operating costs of up to 31%. The results demonstrated that the application of advanced drilling and blasting techniques not only optimizes ore recovery, but also offers a more cost-efficient alternative, even in narrow veins with widths less than 1 meter. The study highlights the feasibility of improving productivity and reducing dilution by utilizing mechanized methods and a more controlled approach in the use of explosives.

On the other hand, the aim of this study [29] was to improve the prediction of blast-induced maximum particle velocity (PPV) in mining, using an optimized random forest model. This model seeks to more accurately predict particle velocity during blasting, a crucial factor in minimizing the undesired effects of explosions on and around mining operations. To achieve this goal, metaheuristic techniques, such as the Whale Optimization Algorithm (WOA), the Gray Wolf Optimization Algorithm (GWO), and the Tunicate Swarm Optimization Algorithm (TSA), were employed to improve the accuracy of the random forest model. These techniques were used to adjust the weights of the decision trees in the model, optimizing blasting parameters such as maximum delay charge and dust factor, which directly influence PPV. The results obtained showed a significant improvement in the accuracy of the model predictions, especially when the optimization algorithms were used, compared to traditional methods. This approach allows us to improve the reliability of PPV predictions, which is essential for the planning and control of blasting in underground mining, contributing to the safety of operations and the reduction of environmental impacts. The study demonstrated that the combination of prediction models with metaheuristic techniques can considerably improve the estimation of the velocity of particles induced by explosions in mining, providing a more accurate and efficient tool to optimize blasting operations in the mining sector.

Likewise, the aim of this study was to identify and characterize the geological and mining factors that determine internal dilution (ID) in lateritic nickel and cobalt deposits in Cuba, to optimize their separation during the extraction process [30]. The study analyzed geological and geochemical data, considering the geological complexity and the mining methods used, to improve the quality of the extracted ore. The methodology used included the observation of the mining fronts and drill holes or boreholes, and the interpretation of the data obtained. In addition, advanced mining technologies were used to evaluate low-grade interlayers and their impact on internal dilution. The analysis detailed the physical, geochemical and geometric properties of non-industrial interlayers, allowing the identification and separation of interlayers immediately before the extracted ore enters the industrial process. The results showed that internal dilution was strongly influenced by geological factors, such as the strength and extent of the interlayers, as well as by field supervision, which plays an essential role in improving mining selectivity. Finally, the study concluded that, with improved knowledge of geological models and the application of more effective control during extraction, the impact of internal dilution can be significantly reduced, thereby improving the efficiency and profitability of nickel and cobalt mining in Cuba.

On the other hand, Delentas et al. [31] proposed to combine empirical approaches and numerical simulations to analyze stability conditions and extracted ore dilution in underground mining by using open stope methods. The work aims to provide easy-to-use mathematical tools to estimate the ore dilution rate and to correlate the mine stability conditions with the characteristics of the extracted ore. To do so, they employed a parametric analysis using RS2 finite element software to model different geomechanically and design conditions of the stope. In their methodology, the stability of the stope, over breaks and the external dilution rate that could arise from failures in the walls of the stope were analyzed. The results of the numerical models were compared with empirical approximations of the stability graphs and mathematical equations were provided to estimate the external dilution rate as a function of basic parameters such as the stability number (N) and the hydraulic radius (HR). The results demonstrated that dilution rates calculated from the finite element model were useful for preliminary stope design, allowing engineers to identify potential dilution issues prior to exploitation. The research also established a relationship between the geotechnical design of the stopes and the stability of the surfaces, with special emphasis on the areas of the sidewalls and the top of the stope. This study offers a combined approach that integrates empirical analysis and numerical simulations to optimize the design of open stopes, reducing ore dilution and increasing the profitability of underground mining operations. The tools developed in the article can be applied for mining project planning, improving operational efficiency and the quality of the extracted ore.

In addition, according to the study conducted by Câmara and Peroni [32], a methodology was developed to quantify the dilution caused by operational efficiency in open-pit mining, specifically in relation to adjacent ore blocks that are not mined correctly due to deficiencies in the operation. The study seeks to improve the accuracy in mine planning by identifying and calculating the dilution resulting from the inability of the equipment to perfectly separate the blocks during extraction. The methodology proposed in the article allows to estimate the dilution caused by operational inefficiency by considering several factors, such as the equipment's ability to remove the blocks, the geometry of the deposits, and the interaction between the ore blocks and the waste blocks. In the process, it is determined which blocks are in contact with the blocks planned to be mined, thus calculating the dilution based on operational deficiencies, such as the lack of selectivity of the equipment or the skill of the operator. This calculation is fundamental to improve the estimates of tonnage and ore grade during the mine planning process. The results obtained demonstrate that dilution can be significant, even when the planning and execution reconciliation process is relatively accurate. The authors concluded that the application of this methodology allows us to identify the causes of dilution and provides a more systematic way of dealing with this problem, which can improve the accuracy of ore resources and reserve calculations, as well as optimize dilution control practices and improve the profitability of the mining operation.

The objective of this research is to explore the impact of using DECKs (inert material placed between the explosive charges) in optimizing the operating charge, to significantly reduce ore dilution and improve rock mass stability. This approach not only promises to improve the structural stability of narrow veins but also represents an efficient and cost-effective solution compared to traditional blasting methods.

The main innovation of this study lies in the use of DECKs instead of conventional solutions such as explosive reduction or blast sequencing. The use of DECKs allows for more precise control of the explosive charge, resulting in a reduction of unwanted fracturing and therefore less dilution. Through simulations and field tests, the impact of this technique on improving safety factors, reducing unwanted fractures and optimizing ore recovery will be assessed. This study will provide new insights into how operating charge optimization with DECKs can transform efficiency in underground mining operations, especially in narrow veins.

The implementation of DECKs not only represents an innovative approach to optimizing operating charge in underground mining but also opens new possibilities for research and application of advanced techniques in blast design. This research seeks to demonstrate that the use of DECKs can significantly improve rock mass stability and recovery, enhancing the profitability and sustainability of underground mining operations.

3.1 Results of geomechanical parameters of drilling and blasting

Geomechanical analysis is a fundamental component of rock mass characterization, as it provides the basis for understanding its behavior under the loads and operating conditions under which mining activity will take place. In this study, the Pampeñita Vein, one of the main geological structures in the Sierra Antapite mining unit, presents parameters including relative density (ρ_r), rock classification using RMR, the GSI index, the rock quality index (RQD), compressive strength (σ_c), and the geomechanical index (Q). These parameters were evaluated for various sections of the vein, including the Caja Piso, Caja Techo, and Mineral areas, with the aim of providing a comprehensive view of the geomechanical variations throughout the structure. The results obtained indicate that the Pampeñita Vein has a fair to good geomechanical quality classification, implying that, although the rock is suitable for mining operations, there are certain areas that may require additional reinforcement due to its degree of fracturing and fragmentation. These values provide key information for the design of support systems and exploitation strategies, aiming to optimize safety and efficiency in drilling and extraction activities.

Detailed results of the geomechanical parameters are presented in Tables 3 and 4.

Table 3. Geomechanical classification of the Pampeñita Vein

Vein	Range/Quality	CP	Cashier Floor	Mineral	CT	Roof Box
		Remote	Immediate		Immediate	Remote
Pampeñita	RMR	56–70	52–65	54–58	45–55	56–70
	Quality (type)	IIIA-II	IIIA	IIIA	IIIB-IIIA	IIIA-II

Table 4. Geomechanical parameters of the Pampeñita Vein

Parameter	Value	Analysis
(Relative density)	2.6	Relative density of the rock, representing a moderately dense material. Important for load and bearing capacity assessments.
RMR (Geomechanical Classification)	56	An RMR rating of 56 indicates a rock mass of fair quality (III-A), suggesting that the ground has moderate resistance to fracturing.
GSI (Geomechanical Index)	51	A GSI of 51 suggests a rock mass with significant fracturing. The quality of the rock mass could be affected by a high degree of fracturing.
RQD (Rock Quality Index)	60%	An RQD of 60% indicates a rock with moderate fragmentation, which may pose a stability challenge in some areas of the mine.
(Compressive Strength)	143	A compressive strength of 143 MPa suggests a high-strength material, suitable for supporting heavy structures without significant deformation.
Q (Geomechanical Index)	3.79	A Q value of 3.79 suggests favorable conditions for stability, but reinforcement may be required in more fractured areas.

The geomechanical classification of the Pampeñita Vein reveals variable rock mass quality, with some areas showing high fragmentation and low cohesion that may compromise stability. This variability highlights the need to adapt support design, with DECKs being an effective solution to improve stability and safety in the most critical zones.

Table 4 shows that the rock mass has generally favorable conditions, but moderate fracturing requires proper support design to ensure stability and safety in mining operations.

3.2 Slot drilling mesh design (free face) and production results

3.2.1 Slot perforation mesh (free face)

The drill pattern designed for the Pampeñita Vein aimed to efficiently generate a slot (free face) to initiate ore extraction. Practical and perimeter-based methods were applied to determine the optimal number of drill holes, considering the geomechanical conditions of the rock mass. The final design, with a 1.50 m × 1.50 m section, included 12-meter-longhole stoping drilled directly into the quartz vein hosting the gold oreization. Production holes were 64 mm in diameter, with reaming at 127 mm. This configuration ensured drilling stability and improved blasting efficiency. The setup is illustrated in Figure 1.

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Figure 1. Equivalent diameter starter design

Application of the Holmberg and Pearson formula:

• Initial parameters

$\emptyset_{{prod}}$= 0.064 m, Production drill diameter

$\emptyset_{{rimado}}$= 0.127 m, Diameter of reamed holes

From Figure 1, the starting design is cylindrical, where D represents the diameter of the equivalent (reamed) drill, yellow, and B represents the burden, distance from the loaded drill (red) to the reamed drill.

The equivalent diameter (D_eq) must be large in relation to the depth of the drill, allowing at least 95% progress in each shot.

3.2.2 Design coefficients

• Burden coefficient for relief drills (K_b) 

When analyzing the stability and performance of relief drill holes in drilling projects, one of the key factors to consider is the burden coefficient (K_b), which varies depending on the hardness of the rock being drilled. This coefficient is essential for determining the efficiency and effectiveness of the drilling process, as it directly influences the selection of equipment and techniques, and impacts operational safety.

The classification of burden coefficients (K_b) according to the type of rock and its hardness is presented in Table 2, which facilitates the identification of the appropriate values for each specific situation in the drilling of relief holes.

Table 5 presents the classification of burden coefficients (K_b) by rock type and hardness, which facilitates the identification of appropriate values for each specific relief drilling situation.

From Table 5, for medium hardness rocks the value of K_b is adjusted from (3 to 4) for narrow veins.

Table 5. Burden coefficients K_b

Rock Type	Hardness	Kb Value
Very soft rock	Low	11-12
Soft rock	Medium-Low	12-15
Medium-hard rock	Medium	15-20
Hard rock	Medium-High	20-25
Very hard rock	High	25-30

• Spacing Coefficient K_s

In the context of drilling, another key parameter to consider is the spacing coefficient (K_s), which is used to determine the appropriate spacing between drill holes. This coefficient varies according to rock hardness, since the strength and fracturability of the material directly influence stress distribution during drilling. Table 6 shows the recommended values for the spacing coefficient (K_s), classified according to rock type and hardness.

Table 6. Spacing coefficient K_s

Rock Type	Hardness	Ks Value
Very soft rock	Low	0.8-1.0
Soft rock	Medium-Low	1.0-1.1
Medium-hard rock	Medium	1.1-1.2
Hard rock	Medium-High	1.2-1.4
Very hard rock	High	1.4-1.6

Table 6 indicates that as rock hardness increases, greater spacing between drill holes is required (higher K_S), while in softer rocks, closer spacing is allowed to optimize drilling. K_S = 1.1, typical range for narrow vein control.

For the design, the requested slot must have dimensions of 1.5 m × 1.5 m, and the equivalent diameter (D_eq) must be determined to obtain homogeneous values in the application of burden and spacing.

$D_{e q}=D_{{rhymed}} \times \sqrt{N}$                      (3)

where,

D_eq: Equivalent diameter

D_rhymed: Reamed drill diameter

N: Total number of reamed drills

The EXSA manual tells us that to generate an optimal free face for a 10 or 12-meter slot, three or four reamed holes should be used. In this sense, four holes are considered.

$D_{e q}=0.127 \times \sqrt{4}=0.254 \mathrm{~m}$

• Calculating the burden

$B=K_b \times D_{e q}$                     (4)

$B=3 \times 0.254=0.76 \mathrm{~m}$

• Spacing calculation

Using Holmberg's formula for spacing and a coefficient of 1.1, common for regular hardness rock.

$S=K_s * B$                    (5)

$S=1.1 \times 0.76=0.84 \mathrm{~m}$

• Calculation of the number of relief holes (N)

To cover the area of the slot, it is necessary to calculate the number of relief holes.

$N=\frac{\text {Total Area}}{B \times S}$                      (6)

Total Area $=1.5 \mathrm{~m} \times 1.5 \mathrm{~m}=2.25 \mathrm{~m}^2$

$N=\frac{2.25 \mathrm{~m}^2}{0.76 \mathrm{~m} \times 0.84 \mathrm{~m}} 3.5 \cong 4$ drills

Rounded to 4, the value of 4 relief holes is within the acceptable range. The value given by EXSA and the Holmberg and Pearson method are checked, and they coincide.

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Figure 2. Slot START design

Figure 2 shows the start with a breaker, which is the center drill, i.e., the initiator for generating free faces. This step is key in mining drilling planning, as it defines the initial geometry of the excavation and establishes the pattern for subsequent work progress. Furthermore, by designing the drill pattern, as illustrated in Figure 3, it is possible to optimize the performance of the MUKI LHBP 2R Jumbo rig, ensuring proper hole distribution, which improves material fragmentation and operational safety.

Figure 3 shows a 1.5 m × 1.5 m cross-section pattern for creating the free face. Four relief holes with a diameter of 127 mm and 13 production holes with a diameter of 64 mm are shown.

3.png

Figure 3. Slot perforation mesh 1.5 m × 1.5 m

A well-designed drilling pattern also facilitates the reduction of operating costs, as it maximizes the efficiency of resource and equipment utilization, minimizing possible failures or deviations during the drilling process.

3.2.3 Production drilling mesh

The Pampeñita Vein has an average thickness of 0.60 m, with a minimum of 0.40 m and a maximum of 1.0 m. Once the free face has been generated, the pit is exploited; for this, the burden and spacing are calculated.

• Calculation of burden and spacing with the Torbica-Lapcevic and Konya Models.

We will compare two models for calculating the load: the modern model and the classic model.

Modern model

$B=0.17 \times P D_e \times\left(\frac{\emptyset_{t a l}}{2 \times K \times \sigma_\tau}\right)$                   (7)

$K=\frac{1-v}{(1+v)(1-2 v)}$                     (8)

where,

$\begin{gathered}B={Burden}(\mathrm{m}) \\ P D_e=51 \mathrm{KBar}=5100 \mathrm{MPa} \text {(Anfo)} \\ \emptyset_{{tal}}=51 \mathrm{~mm}=0.051 \mathrm{~m} ; \sigma_\tau=15 \mathrm{MPa} \\ \sigma_c=150 \mathrm{MPa} ; v=0.17\end{gathered}$

So

$K=\frac{1-0.17}{(1+0.17)(1-2 \times 0.17)}=1.075$

$B=0.17 \times 5100 \mathrm{MPa} \times\left(\frac{0.051 \mathrm{~m}}{2 \times 1.075 \times 15 \mathrm{MPa}}\right)=1.30$

Konya model

$B=0.012 \times\left(2 \times \frac{\rho_e}{\rho_r}+1.5\right) \times \emptyset_{t a l}$                   (9)

where,

$\begin{gathered}B={Burden}(\mathrm{m}) ; \rho_e=0.8\left(\frac{g}{c c}\right) \\ \rho_r=2.6\left(\frac{g}{c c}\right) ; \emptyset_{t a l}=51(\mathrm{~mm})\end{gathered}$

So

$B=0.0 .12 \times\left(2 \times \frac{0.8}{2.6}+1.5\right) \times 51=1.20 \mathrm{~m}$

In burden's calculation the following is observed:

B_max = 1.30

B_mín = 1.20

For the mesh design it is considered:

Burden (Theoretical) = 1.20 m;

Burden (Practical) = 0.60 m;

Spacing = 0.40 m – 0.50 m

A 12 cm deviation will be considered on a 12 m bench (Muki LHBP 2R rig, worn).

Figure 4 shows the drill pattern design for the production holes. This design is optimized for veins less than 1.0 m thick, where a zigzag pattern will be used, and the relief holes will be oriented toward the roof box. This configuration ensures greater blasting efficiency and control, maximizing operational safety and performance. This design not only improves extraction efficiency but also contributes to process safety by minimizing potential ground failures during blasting.

4.png

5.png

Figure 5. Operating load calculation

Figure 5 shows the operating load of 0.50 meters, which is 1.55 kg/delay. The critical speed is 2030 mm/s.

To detail the calculations, it is considered: d = 0.50 m, from the vein to the contact zone:

$K_{{prom}}=357 ; \propto_{{prom}}=-2.07$

$V P P=k \times\left(\frac{d}{Q^{'\frac{1}{3}}}\right)^{-\alpha}$                        (15)

Equalizing

$V P P=V_{ {crit}} ; V_{{crit}}=k \times\left(\frac{d}{Q^{\prime \frac{1}{3}}}\right)^{-\infty}$

Clearing the Q' (Operant Load)

$Q^{\prime}=\left(\frac{V_{{crit}} \times d^{-\alpha}}{k}\right)^{3 /-\alpha}$                     (16)

$Q^{\prime}=\left(\frac{2030 \mathrm{~mm} / \mathrm{seg} \times(0.5)^{-(-2.07)}}{357}\right)^{3 /-(-2.07)}=1.55 \mathrm{~kg}$

• Calculation of the maximum operating load (Q’max).

Figure 6 represents the damage criteria. At 0.50 meters, the critical velocity is 2842 mm/s. This is colored yellow, indicating a damage threshold, with the creation of new fractures.

6.png

Figure 6. Maximum operating load calculation

Damage criterion: (V_crit*1.4) new fractures are generated:

$\begin{gathered}{Vcrit.}=1.4 \times 2030=2842 \frac{\mathrm{~mm}}{\mathrm{seg}} \\ Q^{\prime} \max =\left(\frac{2842 \mathrm{~mm} / \mathrm{seg} \times(0.5)^{-(-2.07)}}{357}\right)^{3 /-(-2.07)}=2.53 \mathrm{Kg}\end{gathered}$

Maximum Operant Load. Load that should not be exceeded according to the vibration limits, the maximum operating load is shown in Figure 7.

7.png

8.png

Figure 8. Maximum operating load

Figure 8 shows the exit sequence from the inside out, starting with number 1, known as the mouth breaker, to generate the free face with the relief holes. Four fanels are used per hole, i.e., four primers.

For a total of 28 fanels, the explosive loading design for the free face is shown in Figure 9.

From Figure 9, the explosive charge design for a 12 m length consists of four primers per hole with the same delay number, starting the exit through the mouth breaker to generate the free face with the reamed holes. The figure was developed by the company using DESWIK.UGDB software.

9.png

Figure 9. Design of explosive charge for the free or slot face with DESWIK software

10.png

Figure 10. Slot damage analysis and simulation (Excel)

Figure 10 shows the simulation of damage in the slot blast.

3.4.2 Blasting design of production drills

A drilling pattern was designed to optimize fragmentation and minimize damage to the rock mass. Appropriate spacing between drill holes was used to ensure efficient blast coverage, as shown in Figure 11.

11.png

Figure 11. Explosive charge design – production drills

The use of DECKs made of dry quartz sand and crushed andesitic gravel was evaluated for their mechanical stability and chemical inertness. Inserted between explosive segments, they enabled controlled energy distribution, reduced unwanted fracturing, improved blast selectivity, and enhanced ore recovery. Both materials were physically and chemically characterized to confirm their effectiveness as energy-dissipating elements. Table 7 presents their key properties.

Table 7. Physical and chemical properties of the inert materials used as DECKs

Property	Quartz-Based Sand	Andesitic Gravel
Particle size	0.3–0.6 mm	6–12 mm
Density (g/cm³)	2.65	2.7
Main composition	>95% silicon dioxide (SiO₂)	Silica and andesite
Water absorption (%)	< 0.5	0.5–1.5
Chemical reactivity	Inert	Inert
Mechanical behavior	Stable and incompressible	High resistance and stability

Figure 11 shows the explosive charge design for production drill holes in narrow veins with a length of 12 meters. Different delays are used, starting from the free face, and DECKs (spacers) are used to fragment the explosive charge.

Figure 12 shows the charge design and output sequence using JK Simblast 2D Ring software, which allows for design, simulation, and testing before blasting.

12.png

Figure 12. Load design and output sequence with JK Simblast 2D Ring Software

Figure 12 also shows the loading design in the drill hole facing the floor box, and an empty drill hole facing the roof box, which is where the greatest control is required. In the loaded drill hole, the ANFO charge is colored yellow, the spacer deck is colored lead, and the primer, which is differentiated by the exit time, is colored red.

Figure 13 shows the blasting simulation.

13.png

Figure 13. Simulation of the influence of explosive energy with JK Simblast

Figure 13 shows the influence of the energy of the explosive as a function of the Load Factor in kg/ton, that is, based on the amount of explosive used for 1 (one) ton of ore. The rock density (quartz vein) is 2.6 g/cc. The simulation shows different colors that indicate the energy levels, the most critical being red, yellow for controlled damage, green where the damage is minimal, and blue for the area where there is no damage.

On the other hand, Figure 14 is presented, which responds to the Simulation of the Influence of the energy of the explosive with JK Simblast.

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Figure 14. Simulation of the influence of explosive energy with JK Simblast

Figure 14 shows the detonation and vibration levels produced by the simulated blast. The sequential detonation of the primers can be seen, starting from the bottom, with the lowest number, 25 milliseconds, followed by the delays spaced 25 ms apart. The colors are damaged indicators: red indicates the crushed zone, yellow indicates the area where new fractures are generated, green indicates areas where pre-existing fractures are present, and blue indicates the unaltered zone.

3.4.3 Comparison before and after the implementation of DECKs

Initially, longhole stoping were fully loaded. After observing the negative results, the use of desks was proposed to reduce the operating load and thus obtain better results.

• Comparison of load design before and after using DECKs.

Figure 15 shows the difference in loading design. The first involves fully loaded holes, that is, to their entire length and with the same number of holes (delays). This generates excess energy during blasting, which directly affects the stability of the rock mass and consequently increases dilution, resulting in ore loss in the pit. The second drill uses DECK technologies, thus reducing the explosive charge, which was previously calculated using the critical velocity. For better control, delays are used with a difference of 25 milliseconds (MS), which reduces the energy generated by the detonation. 

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Figure 15. Load design comparison before and after using DECKs

Figure 15 shows the difference in the charge design, where Emulnor 3000 1 ½ × 12" primers are used, and ANFO as the explosive agent. The charge design on the left side is a traditional charge design, having the same number of delays (25 MS), the primers detonate at the same time, which generates excess energy. For better control of this excess, DECKs are used, their charge design is the figure on the right, when using DECKs, and delays with different numbers, the energy decreases considerably.

• Comparison of results before and after using DECKs.

Figure 16 clearly contrasts the results. The image on the left is the result of a traditional blast, showing a slump, where mineral loss is due to the inability to recover the mineral buried in waste rock banks caused by the excess energy of the explosive. The image on the right is the result of blasting using DECKs, showing good control of the slumps, which favors mineral recovery.

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Figure 16. Results before and after implementation of DECKs

• Design and loading parameters before and after using DECKs.

To clearly visualize the design and loading parameters of longhole stoping in narrow veins, Tables 8 and 9 show all the differences between traditional blasting and another using DECKs.

Table 8. Design parameters without DECKs

Table 8 shows the design and loading parameters for a traditional blast, i.e., one without DECKs. This table shows a PF of 0.6 kg/ton, a very high value due to the high explosive consumption for an ore tonnage of 37 tons/ton. This indicates that the amount of explosive must be reduced, as this excess energy is generating instability, resulting in flaking and loss of ore recovery.

Table 9. Loading parameters without DECKs

Table 9 details the materials required to carry out the blasting drill. The high explosive consumption is notable, at 98.9 kilograms; this amount represents four bags of ANFO. Loading is carried out for all drilled holes; in this case, there are five holes, with four primers per hole. Similarly, there are four fanels, all of which have the same delay.

Table 10. Design Parameters using DECKs

Table 10 outlines the design and loading parameters for a controlled blast using DECKs. This table shows a PF of 0.4 kg/ton, an adequate value with low explosive consumption for an ore tonnage of 37 tons/ton. This indicates that the PF is within the parameters. Explosive energy is controlled, minimizing vibration and consequently controlling rock mass stability. Additionally, accessory consumption and total explosive consumption for detonating four production holes are shown.

Table 11. Loading parameters with DECKs

Table 11 specifies the materials required to carry out the blasting drill. The explosive consumption is notable, at 55.2 kilograms; this amount represents a little more than two bags of ANFO. Loading is not required for all drilled holes; in this case, six holes will be drilled, but only four holes will be loaded, with four primers per hole and using the DECKs. Similarly, the number of fanels is four, all with the same delay.

3.4.4 Comparison of KPIs before and after using DECKs

When viewing Figure 17, the differences existing when using DECKs in blasting longhole stoping in narrow veins are shown. For a bench of 12 meters on average, with vein powers of 0.8 meters on average, with a burden of 1.2 meters; a high contrast can be seen when performing conventional blasting and controlled blasting with DECKs. The number of tons removed is similar in both cases since it is the same pit. This is 37.1 Metric Tons per hole (MT/Tal.), equivalent to one and a half of 23 MT dump trucks.

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Figure 17. Comparison of KPIs before and after using DECKs

Figure 17 shows the KPI comparison before and after using DECKs. The linear charge density when blasting without DECKs is 1.72 kg/m², and the charge when using DECKs is 1.2 kg/m², presenting a difference of 0.52 kg/m². This excess energy is the cause of the instability of the rock mass in the pit. Explosive consumption when using DECKs is 15.3 kg/m², while when not using DECKs, it is 21.34 kg/m². There is a considerable difference in explosive consumption. Using DECKs minimizes explosive costs, which is beneficial for the company. Finally, the efficiency indicator for blasting without DECKs is high, with a PF of 0.62 kg/m², which is considered high for this method and outside the efficiency parameters. On the contrary, when using DECKs, a PF of 0.4 kg/TM is obtained, which is within the appropriate parameters, indicating good performance of the explosive.

3.4.5 Dilution comparison

Mineral dilution refers to the incorporation of unmineralized (or lower value) material into the mined ore, which can lower the average grade of the processed ore and affect the profitability of the mining project. Several strategies have been adopted to mitigate dilution. These include improving the precise design of ore block boundaries, ongoing personnel training to ensure that operators are well-trained, constantly monitoring, and most importantly, improvements in the drilling and blasting techniques that are the subject of this study. Several methods exist for numerically calculating dilution.

$\begin{gathered}{Dilution}=\text {Ton.Stripping/Ton. ore}\left({ }^*\right) \\ {Dilution}=\text {T.Stripping} /({T.Stripping}+  {T.mineral})\left({ }^{* *}\right)\end{gathered}$

The standard of measurement used to measure dilution is the second equation, since it is more sensitive to the increase in stripping.

Figure 18 shows that, with the same pit parameters, the dilution percentage is very high compared to traditional blasting, at 57.14%. This indicates ore loss in the pit due to ore shedding. In contrast, when controlled blasting using DECKs technologies, the dilution is within the planned range of 5.26%, while the planned dilution is 12%. This clearly shows the difference between blasting without DECKs and blasting using DECKs.

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Figure 18. Comparison of dilution before and after using DECKs in blasting

3.4.6 Statistical validation of results

To assess the significance of the observed improvements in ore dilution and rock mass stability after implementing DECKs, statistical analyses were conducted. A two-sample t-test was applied to compare key performance indicators (KPIs), such as ore dilution (%) and PF (kg/TM), between blasts with and without DECKs.

The results show a statistically significant reduction in ore dilution (from 57.14% to 5.26%, p < 0.01), indicating that the improvements are not due to random variation. Similarly, the reduction in PF from 0.6 to 0.4 kg/TM yielded a significant difference (p < 0.05). These findings confirm that the implementation of DECKs has a measurable and statistically supported impact on reducing dilution and optimizing explosive efficiency.

All statistical tests were performed using standard significance thresholds (p < 0.05) and validated with the Shapiro-Wilk test to confirm normal distribution of the data.