Overview Video

1.3.1: Where Are All of These Mines? Continued

Next, let’s look at the fuel minerals, and specifically coal. After looking at Figure 1.3.7, examine the USGS map of U.S. coalfields in Figure 1.3.8. Do you see the correspondence between the coalfields and the location of coal mines? There is a lot of detail to be gleaned from this figure, and I suggest you zoom in and take a closer look. Take note of the rank of the coal found in the different coalfields, and also note the relatively few locations in which metallurgical-grade coal is found.

Figure 1.3.7: Active Coal Mining Operations

The National Institute for Occupational Safety and Health [12] (NIOSH)

For accessible version, contact CDC. [6]

Figure 1.3.8.: Coal Fields of the Conterminous United States

Clickable version via USGS: US Geological Survey [13]. Download PDF for higher quality image. [14]

These figures that we’ve just examined are helpful as we endeavor to learn major characteristics of the U.S. mining industry. There are many ways to “slice and dice” the available data, depending on your specific interests.

1.3.2: Where Are All of These Mines? Continued

In this lesson, I want you to look at a few additional representations. The Mineral Commodity Summaries [15] are published on an annual basis.  This report is the earliest Government publication to furnish estimates covering nonfuel mineral industry data. This publication will break out in more detail the major metal and nonmetal (industrial mineral) mining regions by commodity. Spend some quality time reading and analyzing the Mineral Commodity Summaries to view the figures of the relative economic value or significance of mining in specific regions. The economic value of metals mined in the U.S. is shown by region. We can examine industrial minerals in the same fashion, starting this time with the economic value of industrial minerals production by state. We can then go to a more detailed commodity map to see not only the location at which the major industrial minerals are mined, but also which ones are contributing to the economic value within a given state or region.  

Finally, I would close this discussion that has focused on the location of mining activity with a different perspective on the location of mines. We now understand the geographic dispersion of mining activity as well as the concentration of mining in certain areas for certain commodities. A different twist to the question, where are these mines?, is: are these mines located on the surface or are they deep beneath the surface? Figures 1.3.9 and 1.3.10 clearly show that most mining occurs in surface rather than underground mines, by a ratio of nearly 15:1. We’ll examine the reasons for this later in the course, but an important point now is that most of these mines are located on the surface.

Figure 1.3.9. Active Underground Mining Operations

The National Institute for Occupational Safety and Health [16] (NIOSH)

For accessible version, contact CDC. [6]

Figure 1.3.10. Active Surface Mining Operations

The National Institute for Occupational Safety and Health [17] (NIOSH)

For accessible version, contact CDC. [6]

In this lesson we’ve learned where the fuel and nonfuel minerals are mined, and we’ve developed a better understanding of the extent of mining activities. Soon, we will be ready to look at the life cycle of mining operations and the engineering associated with each part of the cycle. Prior to doing so, there is one last overarching topic to cover, and that is mining in today’s world.

3.2.1: Step 1 -- Preparing the Data for Estimating the Reserve

As a starting point, you’re likely to have the following:

• a table of the coordinates of each drill hole,
• the drill log for each hole, and
• analytical results from each hole.

Here are examples of each of these.

The table of coordinates may look like this:

Here is a section for a typical drill log. The complete drill log for this hole can be viewed here: Driller’s Log.pdf [24], and you should look at the full log.

The analytical results will come from laboratory studies to determine the aforementioned parameters of interest. Here is an example taken from the lab results for the sample obtained from one drill hole.

Figure 3.2.1: Example of analytical results

SGS North America, Inc.

The complete lab report for this hole can be viewed here: Reserve Estimation.pdf [25]

There may be multiple lab reports. The example here focuses on the chemical characteristics of the coal. In many cases, we'll conduct physical tests on the cores to determine geotechnical parameters, e.g. compressive strength, on the ore as well as the rock around the orebody.

We will want to build a database that contains the parameters of interest for each of the holes. If we are interested in determining the average grade, then our table will begin with two columns: drill hole number and the grade for the sample from that hole. Let's suppose that we have a property with 9 holes:

We want the average grade for the deposit. Is the average grade equal to the arithmetic average, which is 3.33%?

Are the holes spaced uniformly on a grid, like this?

Figure 3.2.2:  Boundary and holes

Credit: J. Kohler, © Penn State University, is licensed under CC BY-NC-SA 4.0 [1]

If so, it will be easy to define an area around each hole and then to say that everything within that area has the same properties as those found in the drill hole. Let's draw a box around hole number 5 to illustrate this. Shortly we'll refer to this area as an area of influence.

Figure 3.2.3:  Boundary showing center hole.

Credit: J. Kohler, © Penn State University, is licensed under CC BY-NC-SA 4.0 [1]

If the area around each hole were identical, then it would seem reasonable to say that the average grade of this orebody equals the arithmetic average of 3.33%. But wait a minute! We've said nothing about the thickness of the orebody at each hole. Assuming the thickness is identical at each hole, then each hole will represent an identical volume of ore, and computing the arithmetic average yields the correct average grade for the orebody.  

However, it's rare that the orebody thickness would be the same at each hole. For the purposes of this example, let's assume a more realistic case in which the thicknesses vary from hole-to-hole. Logically then, a hole through a thicker section of the orebody will represent a greater volume of ore than a hole through a thinner section. If we simply average the two holes together, we will arrive at an incorrect average grade because we have not accounted for the larger contribution of the one hole into the total. We can correct this by using a weighted average, in which the grade of the hole is increased or decreased to reflect the volume that it represents.

3.2.2: The Concept of an Area or Volume of Influence

Procedurally, we do this by calculating an area and volume of influence for each hole. The weighted average for the grade, or whatever characteristic is of interest, is obtained by multiplying the value of that characteristic by the volume of influence; and then summing the products and dividing the sum by the sum of the weighted volumes.

Mathematically, this is expressed as:

where $Gn$ = the % grade for the nth hole, and $Vn$ = volume of influence for the nth hole.

Let’s continue with the example by adding the thickness at each hole and inserting columns for the calculated values. The area of influence is the area surrounding each hole, and if the holes are spaced at 400’ intervals, then the area represented by each hole is 1.6x 105ft².

The average grade of the orebody is the weighted grade, 2801 x 10⁶ %-ft³ divided by the volume of influence, 800 x 10⁶ ft³, which equals 3.5%.

Note that the average grade is NOT the arithmetic average of 3.33%. A tenth of a percent error in the grade is quite meaningful. It is important to calculate weighted rather than arithmetic averages in all cases.

In the foregoing example, we had a convenient simplification: the area of influence was the same for each of the holes. In practice, this would rarely be true because the property boundaries are generally irregular and the holes are most likely not spaced evenly. In these common situations, we need a way to determine the influence that a given hole should have in our estimation of the reserve.

Defining an Area or Volume of Influence for a Drill Hole

Consider the following property.

Figure 3.2.4: Property with 12 holes.

Credit: J. Kohler, © Penn State University, is licensed under CC BY-NC-SA 4.0 [1]

The new challenge here is to determine the area of influence for each hole. Once we have done that, we can continue by using the same procedure that we followed for the previous example.

Each hole is likely to have a different value for the characteristics of interest, and for this discussion let’s say that we are looking at grade. How far from the hole should we assume that the grade of that hole applies? Halfway to an adjacent hole? What if the grade in the adjacent hole is significantly different? Should that alter where we draw the area of influence? Perhaps, we should use a scheme that says the value at the hole decreases inversely as we go further from the hole? In fact, many deterministic and statistical methods have been developed over the years, and some provide better results than others for certain types of ore bodies. Let's take a look at a few methods for determining an area of influence for each drill hole.

3.2.3: Overview of Reserve Estimation Methods

The polygon method is an old and established approach based on a simple geometric algorithm, in which we construct a polygon around each hole to determine an area of influence for that hole; and then the total volume directly beneath the polygon is assigned the same values as the drill hole from which we constructed the polygon. We’ll take a closer look at his method shortly.

Another method, known as the triangle method, requires that we connect adjacent holes into triangles. The included area of each triangle is assigned the characteristic not of a single hole, but of the weighted average of the three holes forming the triangle. The weighting of the three holes is based on the length of the drill holes.

The inverse distance method is a more complex scheme in which the contribution of a given hole is weighted according to its distance from the block in which the estimate is to be made. The closer a hole is, the more weight is given to its value compared to the values of other holes in the region.

Geostatistical methods employed for ore reserve estimation utilize three-dimensional spatial statistics to improve the quality of the estimate. Classical statistics requires use of a particular distribution model, e.g., the data are normally distributed, and that the samples be independent of one another. Generally, we have insufficient samples of the orebody to assign a distribution, and moreover, the samples are often correlated, i.e., they do not satisfy the independence requirement of classical statistics. Given that the samples, i.e., the drillholes, are limited in number because they are expensive to acquire, often biased, and nearly always smaller in number than is desired, geostatistics is a powerful tool for improving the quality of the estimation.

A prerequisite to a reasonable prediction of the grade of the orebody is a good prediction of the spatial distribution of the grade, or whatever characteristic is of interest. This spatial estimation is accomplished using the sample data and a model known as a variogram, which is used to represent the correlation between the samples. This estimation is often accomplished using a technique known as kriging. Kriging provides an optimal interpolation using the variogram; and the technique is similar to simple interpolation, as we would use in say the inverse distance algorithm, but is different, because it allows us to take into account information that we know about the geology and attendant properties. Based on geologic knowledge of the presence of a certain feature, we will know that the characteristics of all points contained in that feature should be the same or similar. This is an instance where samples are correlated. With geostatistical methods, we can use this knowledge to improve the estimate of grade, or whatever, at points where we have not sampled. The science of geostatistics continues to evolve, becoming more accurate. You will learn the common geostatistical techniques, their strengths, and limitations in MNG 412.

Today, you can enter your exploration data into a computer program and then within minutes, you can have estimates of the resource from several different methods. Then, you will choose which estimate to use. That choice may be based on experience or a heuristic, such as selecting the estimate that has the smallest variance. The commercially available mine planning software programs, such as the Carlson software that we use, allow you to employ several different techniques. We are not going to look into these methods in any greater details in this course, with one exception --the polygon method. This method is useful to illustrate the concepts and is a reasonable estimation method in its own right.

The Polygon Method

We begin with a map showing the surface location of the drill holes, and our task is to construct polygons around each hole. The relevant characteristic, say grade, inside of that entire polygon will be the same as the value of that characteristic in the drill hole sample.

We start our work by arbitrarily selecting a drill hole, and then drawing lines between that hole and all the adjacent holes, as shown here.

Figure 3.2.5: Left (1): drill hole, and Right: (2) lines drawn between the hole and all adjacent holes

From p. 78 of Introductory Mining Engineering, 2nd Ed., Hartman, H.L, and J.M. Mutmansky

Next, we draw perpendicular bisectors though each of these lines, drawing the bisector line long enough to intersect the other perpendicular bisectors, as shown.

Figure 3.2.6: Bisector lines are long enough to intersect other perpendicular bisectors

From p.78 of Introductory Mining Engineering, 2nd Ed., Hartman, H.L, and J.M. Mutmansky

The corners of the polygon are defined by the intersection of the perpendicular bisectors, as shown.

Figure 3.2.7: Polygon with corners defined by perpendicular bisectors

From p. 78 of Introductory Mining Engineering, 2nd Ed., Hartman, H.L, and J.M. Mutmansky

Watch this video (2:59) of a demonstration of using the Polygon Method.

We repeat the process for each hole. Note that if the hole is adjacent to the property boundary, then that boundary line will form a side of the polygon. The result will be a property containing as many polygons as holes, as illustrated here.

Figure 3.2.8: The result is a property with as many polygons as holes

From p. 78 of Introductory Mining Engineering, 2nd Ed., Hartman, H.L, and J.M. Mutmansky

Next, we need to determine the area of each polygon. This can be done manually using a planimeter, or digitally. The result will be an area of influence, i.e., the area of the polygon, for each hole. Then, we can calculate the volume of influence of each polygon, by multiplying the area of influence by the thickness of the ore, or overburden. The next step is to build a table or spreadsheet to facilitate the calculations. We actually did that earlier, in Lesson 3.2.2, and will not repeat it again. At that time, we only calculated the average grade. We could have added any number of other characteristics to the table, and calculated their average value. Examples would include thickness of the deposit and overburden, as well as other characteristics of interest.

3.2.4: Step 2 -- How Much of the Resource is a Reserve?

We started this lesson by noting that reserve estimation is completed in two steps. The first is to estimate the size of the resource, which we have now done. The second step is to determine how much of the resource can be mined economically at a point in time.

We’ll need to learn more about mining methods to tackle that question completely, and we will do so in the coming weeks. Notwithstanding, there are two metrics that can be computed very early in this second step. The first metric is known as the cutoff grade, which is basically the lowest grade that can be mined at a profit. The second metric is applicable to shallower deposits that are being evaluated for surface rather than underground mining, and it is known as the stripping ratio. Let’s start with stripping ratios, and let’s use a shallow coal seam for our example.

Stripping Ratios

The coal seam will be underneath layers of soil and rock. The material overlying the seam is known as the overburden. Before we can extract the coal, we first have to remove, i.e., strip, this overburden. It costs money to remove the overburden, and in the simplest terms, the cost of removing the overburden cannot exceed the value of the coal that is exposed.

The stripping ratio is usually taken as the volume of the overburden that must be removed to the weight of the coal that is exposed when this volume of overburden is removed. Thus, the units for the stripping ratio will be yd3/ton. Two stripping ratios are used in the prefeasibility or feasibility analyses: overall stripping ratio and the maximum stripping ratio. As mine planning advances beyond the prefeasibility stage, stripping ratios at different cross-sections will be calculated as well.

The overall stripping ratio is calculated using the average values for the volume of the overburden and the average value for the weight of the coal (ore). This number is a key indicator for the potential of the project to be profitable. Please remember that these averages are weighted averages.

The maximum stripping ratio, which is also known as the breakeven stripping ratio, is an economic calculation based on the cost of removing the overburden and the value of the coal or ore that is exposed when the overburden is removed. Thus, given the stripping cost and the value of the exposed ore, we can calculate the breakeven or maximum stripping ratio. Stripping costs can be estimated reasonably well, based on the method of overburden removal and the region in which the mining is being conducted. We can find tables of data in handbooks to help us with this estimation. The value of the coal or exposed ore is usually taken to be its selling price. The selling price must account for the mining and processing costs, and the minimum profit that the company requires.

Stripping cost = Value of the exposed coal

where:

• SC = stripping cost, $/yd³
• V_ob= volume of overburden that will be removed at breakeven
• V_c= volume of coal that will be exposed by the removal of Vob
• t = thickness of the coal seam
• TF = tonnage factor (density) of the coal, tons/yd³
• SP = selling price of the coal, $/ton

The breakeven or maximum stripping ratio, SRmax, is, therefore:

The value and use of SR_max is illustrated in two examples. Consider the situation represented in the following figure of a coal stripping operation.

Figure 3.2.9: Coal Stripping Operation

Figure 7.9 from Introductory Mining Engineering, 2nd Ed., Hartman, H.L, and J.M. Mutmansky

The coal seam is under a hill. Most likely, mining started at the edge of the hill where the coal seam outcropped, i.e., intersected the surface. As mining progresses back into the hill, it will be necessary to remove increasing amounts of overburden to expose and mine the coal. At what point does it become uneconomical to remove the overburden to access the coal? Correct –at SR_max! And how do we find SR_max? Correct –using Equation 3.2.3. And therefore, at the point where we are at this maximum stripping ratio, SR_max we will stop mining.

The calculation of stripping ration is slightly more complicated for deposits that are not flat lying, i.e., they are dipping at an angle greater than a few degrees; although, conceptually, the process is the same regardless of the spatial characteristics of the deposit. As you'll see, the math is slightly more involved to complete the calculation of the stripping ratio. Let's take a look. 

3.2.5: Instantaneous Stripping Ratio

The stripping ratio (SR) refers to the amount of waste material that must be removed for a given amount of ore. The Instantaneous Stripping Ratio (ISR) is the stripping ratio for a given push back, where a tiny slice of material, i.e., ore and/or waste, is removed from a pit wall. This section presents the ISR calculation for a steeply pitching deposit.

Assume an idealized tabular and steeply pitching orebody that outcrops at the surface and dips to the left at ϴ degrees (Figure 3.2.10). Assume that the ore extends down to considerable depth, and that open-pit mining will be used to extract ore. When open-pit mining is no longer economical, an underground mining method will be used to recover the ore. Therefore, we will need to calculate the point at which we will cease surface mining and either go underground or close the mine.

The over-lying waste, i.e., the non-mineral-bearing rock, must be removed to uncover and mine the ore. The shape and size of the pit depends upon economic, engineering, and production factors. Assuming all other factors to be constant, as the selling price of the ore increases, the pit size will increase.

Figure 3.2.10: An idealized tabular and steeply pitching orebody

Source: A. Lashgari. © Penn State University, is licensed under CC BY-NC-SA 4.0 [1]. 

If you were going to calculate the ISR for a “real” orebody for which you had drillhole data, you would utilize one of the computer software packages, such as Carlson. Here the purpose is to teach you the principles, so we are going to make some assumptions to simplify the calculations, and to better illuminate the procedure without getting buried in the math. The assumptions are as follows.

• The seam thickness is constant. The orebody thickness is represented by the length of the ore section (Lo) in the previous diagram. Note that the length of the waste section (Lw) increases as our pit advances deeper. Therefore, as we advance, more of the waste will have to be removed to expose a unit of ore, as shown in the next diagram.
• The density of ore and waste are the same.
• All slices have the same strip width (Ws).
• We are looking at a two-diminsional slice of the Earth’s crust. Obviously, the ore and overburden continue into the page and go on for some distance. For our purposes here, we will assume that they go into the page for a distance of 1 unit.
• The slices of the material are mined out perpendicular to the orebody.

Watch this video (3:11) on an explanation of the instantaneous stripping ratio.

Later in this section, I will explain how each of these assumptions affects the calculations. However, it would be instructive for you to pause for a moment and think about each assumption.

Figure 3.2.11:

Source: A. Lashgari © Penn State University, is licensed under CC BY-NC-SA 4.0 [1]

Now, let’s calculate the ISR for strip 1:

where T is strip thickness, L is strip length, W is strip width and $\rho$ is material density.

Since the density, width, and thickness of the ore and waste are assumed to be equal, this equation reduces to:

Assume that the ore seam thickness is 40 ft and the length of the waste slice on top of the ore in strips 1, 2 and 3 is 45, 50 and 55, respectively. The ISR for these three strips is calculated as:

This simplified example has illustrated how the length of waste and ore sections will impact the ISR. In a mine, there will be differences in the density of the ore and waste materials. However, the difference can be trivial in some cases. In fact, a real-world example is where there are several waste types in the slice, and the waste materials have densities that are each different from the ore. Therefore, the calculation of ISR using the length ratio does not work. Instead, a generalized IRS calculation must be defined as:

where there are $i$ different waste material types in the strip.

If the density differences are not drastic, the simplified form of the calculation will give you a rough estimate of the ISR. If you need more accurate values with higher accuracy level, then you should consider all other parameters in your calculations. Of course, as I mentioned earlier, if you are doing a complex and “real-world” case, you will probably be using a software package; and then, accounting for the myriad of details becomes easier.

By the way, please note that the ISR is independent of the deposit's dip angle.

For this discussion, we have used an orebody of uniform shape and plunging at some angle θ. But what of a deposit with an irregular shape?  The same process applies. Therefore, regardless of the shape of the orebody, the above equation can be used to determine ISR for a strip extracted from the side wall of the pit. It should be also noted that different units can be used to express the ISR.

The most popular units for the ISR are tons of waste removed/ton of ore exposed, ft of waste removed/ft of ore exposed, and yd³ of waste removed/ton of ore exposed. The latter is the most common unit in many types of nonmetal operations because there is no value to know the weight of the waste, unless it poses a limit to the trucks. Therefore, the density of waste material does not come into play in the calculations. If there is an issue with a weight limit in the haul trucks, then the tons of waste/ton of ore is an appropriate unit for the ISR or SR.

3.2.6: Maximum Allowable Stripping Ratio (SRmax)

Earlier in this lesson, we looked at the maximum stripping ratio, and we did it for coal seam. You will recall that the maximum allowable stripping ratio, SR_max, also called break-even stripping ratio, is the maximum amount of overburden/waste that can be extracted per unit of ore at the economic pit limit. The SR_max is determined solely by economics, to establish the ultimate boundary of the pit, where break even occurs, i.e. the profit margin is zero. As we defined it before,

So, it is a physical quantity that is determined by economics. This value can be simply converted to the unit of tons_w/ton_o, considering the density of the waste material. If the ISR exceeds the SR_max, then the operation will be uneconomical, as the income generated by the ore will be insufficient to offset the costs incurred in mining.

Now, let’s imagine a massive irregular deposit, where copper ore is the only ore that is desired to be mined out. Unlike coal deposits, metals are not extracted in their native form, except in rare cases. Instead, the rock has a small percentage of valuable minerals in it. A copper deposit contains rock that can be profitably mined and processed to extract the copper. However, the amount of copper contained within the rock, i.e. the grade, varies by location. We’ll need to account for this in our calculation of the break-even stripping ratio.

The only difference is that ore grade variation should be taken into account in the calculation. In order to determine the SR_max for such a deposit, the orebody is divided into different blocks of ore. The average ore grade for each block is determined, and then the overall grade of the ore in the slice is calculated as follows:

where g ¯ is the average ore grade in the slice, L oi is the length of the ore section that has a grade equal to g oi .

After you calculate the average ore grade for the slice, you can use a grade - stripping ratio (g-SR) plot to determine the SR_max associated with the determined average ore grade. Here is an example of a g-SR plot.

Figure 3.2.12: Grade-stripping ratio (g-SR) plot

Source: A. Lashgari © Penn State University, is licensed under CC BY-NC-SA 4.0 [1]

Imagine a copper deposit in which the average copper grade for a strip is 1.05%. Checking the g-SR plot for that deposit, suppose we find out that the SR_max for g ¯ = 1.05 is equal to 8.5. This means that 8.5 units of waste can be economically removed per unit of ore. If the ISR for the strip is smaller than the SR_max, then the pit could be extended and more strips could be mined profitably. If the ISR and SR_max are equal, then this is a good place for the pit limit. If the ISR is larger than SR_max, then we have passed the economic location for the pit limit.

In MNG 441, you will learn how to determine break-even cutoff grade and draw a g-SR plot using economic parameters. Here, I simply want you to know about this plot and what it is representing.

3.2.7: Impact of Pit Expansion on Stripping Ratio

Imagine a steeply pitching orebody, as shown below. We would like to extract this tabular deposit using open pit mining method. Let’s assume that two pits with different sizes are being considered for this mining operation. We would like to study the impact of pit expansion on stripping ratio by comparing the overall stripping ratios for these two pits.

This figure shows the pit areas as a block model. The smaller pit (Pit A) includes 32 blocks. 16 blocks are ore and the other 16 are waste blocks. Therefore, the stripping ratio is 16/16 = 1. The larger pit includes 437 blocks in total. Using Pit 2, 76 blocks of ore can be extracted. Therefore, the stripping ratio is: (437-76)/76 = 4.75. This example shows that as we expand the pit in a steeply pitching orebody, overall stripping ratio dramatically increases. Therefore, more waste blocks have to be mined out to uncover one block of ore. It should be noted that as we go deeper in an open pit, the unit cost of mining will increase. Because the material will be transported for a longer distance, which takes more time, and which may necessitate that we add additional trucks to the fleet.

Figure 3.2.13: A block model of the pit areas

Source: A. Lashgari © Penn State University, is licensed under CC BY-NC-SA 4.0 [1]

Now, you should be able to:

1. Determine ISR for given pit dimensions
2. Find the pit limit when the pit dimensions, average ore grade of the strip and g-SR plot are given.

3.2.8: The Concept of a Cutoff Grade

We know that orebody characteristics such as the grade vary spatially within the orebody, and as we approach the boundaries of the deposit, the grade will begin to decrease to zero. Eventually, the cost to mine and process a ton of ore will exceed the value that we can obtain by selling the commodity found in that ton of ore. Mining companies, like other businesses, do not stay in business by selling their product at a loss!

Thus, when we are estimating the amount of the resource that can be mined economically, we’ll have to calculate a cutoff grade; and any ore with a grade lower than the cutoff cannot be counted in the estimate of the reserve. The cutoff grade is the grade at which the cost of mining and processing the ore is equal to the desired selling price of the commodity extracted from the ore.

The cutoff grade is influenced by a few external factors that you can control to a certain extent, and these will be considered in the prefeasibility analysis when the cutoff grade is determined.

Let’s imagine that we have a deposit of copper that we are evaluating. It is a large resource with a grade of 1%. The block below represents 1 ton (T) of in-place ore. We know that the grade is 1%, so we have 20 lb. of copper distributed in that ton of ore. Let’s represent that 20 pounds of copper as a small (not-to-scale) block inside the larger block.

Figure 3.2.14: Figure 6 from the text

Credit: J. Kohler, © Penn State University, is licensed under CC BY-NC-SA 4.0 [1]

When we mine this ton of ore, it is likely that we will get some dilution, which occurs when we extract some of the rock surrounding the orebody. In other words we are diluting our ore with this host rock. Why would we do that? We don’t do it on purpose, generally, but, it is a consequence of the mining method and equipment that we select to mine the ore. Imagine that you have a chocolate cake, and this cake has a thin band of raspberry filling in the middle of the cake. Suppose that you are tasked with removing the raspberry layer, and that you do not want to have any of the chocolate layers mixed with the raspberry layer that you are extracting. Further, I will give you a tiny, baby-sized spoon to complete the task on your cake, and I will give your classmate a large serving-size scoop to complete the task on his cake. You both have ten minutes to complete the job, with a goal of minimizing the dilution of the raspberry layer with the chocolate cake..

The outcome of this experiment is evident. The use of the tiny, rather than large spoon allows the competitor to be more exacting in the removal of the raspberry layer, resulting in far less contamination, i.e. dilution, of the product. In mining terms, we refer to this as the selectivity of the mining method, and some methods are much more selective than others, resulting in far less dilution. Of course, there are tradeoffs between selectivity and other important metrics, such as mining cost and productivity. We’ll look closely at this when we study the mining methods. Now, we can continue with our example.

We extract one ton of material, and now we know that this one-ton of material will be diluted with non-ore bearing material, based on how selective our mining method is estimated to be. We can represent this as shown in Figure 3.2.15 by including the waste within the one ton of mined ore. This waste material effectively reduces the copper in the block by the amount of the dilution. Let’s assume an average dilution of 5%. The copper present in the mined block of ore is reduced by 5%, to 19 lb of copper.

Figure 3.2.15: Figure 7 from the text

Credit: J. Kohler, © Penn State University, is licensed under CC BY-NC-SA 4.0 [1]

Next, our one-ton block of mined material will be routed into the mineral processing plant, where the ore will be separated and concentrated into a saleable product. There will be two output streams from the plant: one containing the saleable product and the other, the tailings, or waste. The physical and chemical methods used to beneficiate the ore are not perfect, and while we can invest more money to improve them, there is a point of diminishing return. Consequently, some saleable product will report to the waste stream, and as such represents a loss. Let’s assume a plant recovery of 90%, meaning that all but 10% of the product in the plant feed will be recovered and report to the product stream. In our example, there is 19 lb. of copper in the feed to the plant and the plant will recover all but 1.9 lb. Thus, the one ton of material that we mined will yield 17.1 lb of copper for sale.

Figure 3.2.16: Figure 8 from the text

Credit: J. Kohler, © Penn State University, is licensed under CC BY-NC-SA 4.0 [1]

Mathematically, the minimum acceptable grade to mine, i.e. the cutoff grade, occurs when the cost to mine and process the material, (M&P cost), is equal to the value of the product, i.e. its selling price, SP.

We will account for the dilution and recovery losses in the right side of the equation, as these losses are reducing the value of the in-place ore.

where:

• M&P cost = mining & processing cost, $/T
• SP = selling price of the copper, $/lb
• D = dilution
• R = plant recovery
• G = grade

Let’s use some numbers. Assume the following: we have a dilution of 5%, a plant recovery of 92%, a mining and processing cost of $6.80/ton and a selling price of $0.74/lb. Be careful with the units!

Solving the equation for grade, we have:

And substituting our values,

$$\text{G}=\text{(\$6}\text{.80}/\text{T){\$0}\text{.74}/\text{lb}*\text{(1}-\text{0}\text{.05)}*\text{(0}\text{.92)}*\text{(2000lb}/\text{T)}}$$ $$\text{G}=\text{0}\text{.499}\cong \text{0}\text{.5}=\text{.5\%}$$

Therefore for this example, the grade of the ore that we choose to mine must have a grade of 0.5% of greater. In other words, the cutoff grade has been determined to be 0.5%, and when you report the size of your reserve, you will include only the tonnage at or greater than 0.5%. Do you understand why you must report it in that way?

Imagine a situation in which your resource contains significant tonnage at slightly less than 0.5%. Likely you will reexamine your dilution and plant recovery assumptions, and explore options for improving both so that you can economically mine the lower grade ore.

We didn’t say much about the selling price itself. Presumable, it is the maximum price that we believe we can reasonably expect to sell our product into the market. The price will need to cover not only our mining and processing costs, but also whatever profit we require. When companies evaluate the merits of opening a new mine they will have certain criteria for establishing the financial merits of the project. They might, for example, require a certain return on their invested capital, a payback within so many years, and so forth. This can all be bundled, at least conceptually, into the required selling price.

Sometimes we will talk about the gross value of the mined ore. The gross value is equal to:

The unit profit is then defined as: Gross Value – Mining & Processing Cost.

To illustrate, assume that we are mining an orebody with a grade of 1.2%, dilution is negligible, and the selling price is $0.74/lb. The gross value is calculated to be $16.34/T and the unit profit is $9.54.

4.1.1: Locational Factors

Ore bodies are located where you found them, and often are not ideally located by any definition. They may be inside the Arctic Circle, high up in the Andes Mountains of Peru, or in the tropics of Indonesia to illustrate just a few out-of-the-way places. Occasionally, they are near metropolitan areas or small towns. Regardless, you will assess the following:

Transportation Options

You must get your product to market, and you must be able to get supplies to the mine. Transportation options suited to the one need may be unsuitable for the other, so both must be independently evaluated. Generally, access to rail is necessary. Where is the nearest railway, and will they be willing to serve a spur to your property? Can you get a right-of-way to build a spur? A highway suitable for tractor-trailer use may be needed as well, or in lieu of a railhead on your property. Sometimes, it will be appropriate to move your product on a waterway, e.g., a river or even the ocean. Can you access a load-out facility to get your product on barges or boats, or do you need to build one? These costs will have to be estimated at this stage, as the costs to get supplies to the mine can raise your production costs. Similarly, your customers will bear the cost of transporting the product in most cases. Although you may be able to mine at a competitive cost, your transportation costs could adversely affect your competitiveness. If you are producing a construction aggregate, your practical limit for transporting the product is on the order of 50 miles. If you are producing iron ore, you may be able to transport it for hundreds or even thousands of miles.

Availability of Labor

The labor requirements will depend on the size and type of mine, but in general, you will require experienced miners who are familiar with the method, equipment operators, welders, mechanics, electricians, and skilled managers. Electronics and computer technicians, surveyors, engineers, and accountants are normally part of the human resource requirements, as well. Some of these services may be provided on a regional basis to multiple mines, especially if your new mine will be part of a large company with other mines. If you’re evaluating a potential stone quarry in the Atlanta, Georgia area, labor will be readily available. If the proposed project is an underground gold mine in rural Nevada, you may have to bus the workers for approximately two hours from the nearest town to the mine site –and you will do this each day. If the proposed mine is located in the Australian outback, you will probably set up a fly-in fly-out operation, in which you use jet charters to transport your workers. They will then remain on-site for two or three weeks, working every day, and then return to their home city for a week or so of leave prior to beginning the cycle all over again. You get the idea! Labor availability can span these extremes, and be anywhere in between. Regardless, the effort and the cost to staff the proposed operation will be considered early on.

Infrastructure

The mining operation will require infrastructure, i.e., electricity, water, roads, buildings, housing, hospitals, schools, and so on. Is this already there, or will you have to create it? How long will it take, and what will it cost? Mining engineers of the 19th and early 20th century were, by necessity, quite skilled at building towns and the necessary infrastructure, as well as opening and operating mines. Companies operating internationally in remote or underdeveloped locations still find the need today to develop the infrastructure. The population in the region of these mining operations often experiences a significant improvement in the quality of life, e.g., availability of clean water, excellent medical care, good schools, and so on. Again, however, an early decision includes consideration of the time, effort, and cost to establish such an infrastructure.

Employee Satisfaction

The location and its climate can affect employee satisfaction, which will impact productivity of the workers and your ability to recruit and retain the workforce. These can be significant risks for the proposed project. If the mine is located near an urban area with an abundance of cultural and entertainment opportunities, and the climate is pleasant year round, then your employee satisfaction is likely to be very high. On the other hand, if the mine is inside the Arctic Circle, you may face significant challenges recruiting and retaining a workforce. Of course, you can “buy” some measure of employee satisfaction with high wages and large bonuses.

4.1.2: Natural and Geologic Factors

The technical characteristics of the orebody, the geologic setting in which the orebody is found, and the surface features of the land over the orebody will influence significantly your choice of a mining method and the way that you will lay out the mine and supporting infrastructure. The most important natural and geologic factors that will be assessed include:

Topography

The location of your surface facilities, e.g., shops, warehouses, and roads, will be influenced by the topography. Mountainous terrain creates more challenges than a gently rolling countryside.

Spatial Characteristics of the Orebody

The depth, size, shape, and attitude will have a major impact on the type of mine and mining operation that you can have. A shallow orebody allows consideration of a surface rather than underground mine. A large orebody allows for larger-scale and potentially cheaper bulk methods, and a long life means more time to recover certain costs. The orebody may be tabular in shape, which lends itself to certain types of mining, or it may be a big amorphous blob, or perhaps the ore is contained in sinuous veins. These shapes, in conjunction with other characteristics, will suggest certain mining methods. And finally, there is attitude. The attitude of the orebody is important. After all, who wants to develop a mine in an orebody that has a bad attitude!  Just kidding... attitude is a term for the angular orientation of the orebody, and, in particular, we are interested in the vertical angle of the orebody. If the deposit is flat lying, the vertical angle or dip is zero degrees. As we begin to tilt one end of the orebody downward, the dip angle is measured with respect to the horizon. Steeply pitching ore bodies with dip angles of 70 degrees or more are not uncommon. As it turns out, some mining methods require steeply pitching deposits, while others work best with dip angles approaching zero degrees.

Geomechanical Properties of the Orebody and the Surrounding Rock

Undoubtedly, we will want to create openings in the orebody or the surrounding rock as part of the mining process. How difficult will it be, and how much energy will it take to break the ore and the surrounding rock? Will explosives be required? Once we remove a section of ore, will the opening be stable, or will the surrounding ore rush in to fill the void? When the ore is removed, will the surrounding rock remain stable, or will it fracture and cave-in? The answers to these questions depend largely on the mechanical properties of the materials. Compressive strength, modulus of elasticity, hardness, and abrasiveness are some of the important properties that will determine how challenging it will be to safely and productively operate the mine. We can take core samples and conduct laboratory tests to quantify the characteristics, and we can avail ourselves of any data that others have compiled on similar deposits.

Geology

Structures such as cleavage patterns, joint sets, and faults can significantly affect the mineability as well as the stability of the materials, regardless of certain mechanical properties such as strength. The stratigraphy, or layers of rock formations, is important in the design of many mine types. The mineralogy of the deposit and the genesis of the orebody will give us an indication of both mining and subsequent mineral processing challenges. The presence of thermal gradients will likely require expensive cooling of the ventilating air. Water-bearing strata or aquifers will increase the complexity of the mining as we work to protect them as well as to deal with groundwater inflows to the mine. There are other examples that will be considered early in the evaluation process, and this once again illustrates the importance of including geologists on the team.

Ore Beneficiation

The ore has physical and chemical properties that will affect the way in which we extract the valuable components from the mined ore. Some copper ores are easier to process than others based on the mineralization, orebody genesis, and so on. These factors will affect the cost of the mineral processing and will be examined before a decision to move forward with the project is made.

4.1.3: Socioeconomic Factors

This category is intended to capture a disparate group of factors that are often beyond the control of the mining company, and which can occasionally create large problems for the mining company.

Demographics

The population characteristics are especially relevant to the labor force considerations, which we discussed under locational factors.

Financing and Market Conditions

Generally, mining companies will want to raise capital to finance the proposed mining venture. The prevailing condition of the financial markets will impact the ease or difficulty of raising capital. Similarly, prevailing regional, national, and global economic conditions can radically alter the demand for mined products as well as the price paid for those products. Sometimes, these conditions can change radically from when a project was begun and when it came online. In recent years, the simultaneous crash of both the commodity and energy markets has had disastrous effects for a number of mining companies.

Government Actions

Governments at all levels can incentivize mining by giving tax breaks for example, or they can restrict mining by delaying permit approvals, among other actions. The degree of government regulation is a consideration, and especially the consistency of interpretation and enforcement of regulations, as these can affect the cost of compliance.  

Taxes

Tax policies such as income tax and depletion allowances will affect the profitability and, hence, feasibility of a proposed project.

Political Stability

Undertaking projects in lesser-developed countries(LDC) presents a higher risk for investors, and this may make it difficult to obtain financing at a reasonable rate, if at all. Governments change and policies can be overturned or reversed in an instant, and what was a welcoming and investment-friendly government can become quite hostile. Civil wars and terrorism can make it difficult to operate, not to mention the difficulty in recruiting and retaining a skilled workforce. Sometimes, governments will expropriate the company’s assets in the country, e.g., the mine, the equipment, and the reserves – this may occur years after the mine has opened. Despite the risk, companies and investors continue to pursue projects in these areas. Why? The “reward” for investing in higher-risk projects is a much higher rate of return on the investment. Thus, in the prefeasibility stage, the analysis would have to support the likelihood of a much higher rate of return. Otherwise, the project will be a nonstarter. It is difficult to account for these factors, not only because they are largely out of the control of the company, but also because they are difficult to quantify and predict. One can use the historical record as well as seek advice from outside experts, e.g., economists, political analysts in the U.S. State Department, and so on. Practically, we can account for some of these risks by performing risk and sensitivity analyses. For example, if we estimate that the likely selling price is $150/ton, and a worst-case scenario is a selling price of $100/ton, we would evaluate the desirability of the project over this range, and report our findings to those making the decision on the project.

4.1.4: Do You Have an Envelope That I Can Use?

Engineers often talk about doing “back of the envelope” calculations. This is an expression for making a first approximation to a solution or answer. It is a process in which we don’t worry about the details, but only the biggest factors that influence the result. By necessity, it requires experience and judgment. The answer is nothing more than a rough estimate and is likely to result in a solution that looks like 1, 10, or 100, i.e., an order-of-magnitude estimate, rather than an exact number like 0.67, 13, or 180! Or, to be more specific here, the team will evaluate the foregoing factors with the goal of concluding one of three outcomes: this project has little chance of success – we need to cut our losses and move on; this project shows strong potential – we need to move into the next stage of the project without delay; or, this project has potential, but without doing further analysis, it is difficult to say that we should move this project to the next stage. Sure, there are shades of gray, but you get the idea of what is happening at this stage.

Reasonably, you may also wonder just who are the people making this decision. That’s a good question. Although the full team may include a number of subject experts, e.g., geologists, mining engineers, among others, the decision-makers are likely to be few in number and are likely to include: a mining engineer with decades of experience in the industry and years of experience in evaluating and bringing new projects online. This person probably has a title with the words “vice-president” in it. There will be a financial wizard. Someone with years of experience in the industry, who can probably do complex discounted cash flow analyses in her head. And there may be a third – perhaps a “rising star” in the organization. Someone with several years of experience, highly motivated, and with the potential to head up this new mining project if it becomes a reality. The model will vary depending on the size of the company and the size of the project, but this will give you an idea of the approach.

So, what’s next? Assuming that the conclusion of our “back-of-the-envelope” analysis is that we still want to build a mine, then we will move to the feasibility-study. This is the subject of Lesson 4.2.

Lesson 4.1 Summary

The major factors that will affect the desirability and feasibility of a mining project can be divided into three categories. The consideration of the factors will precede any decision to conduct more detailed studies. It is important to understand the three categories, the factors within each category, and in what manner they influence the decision.

4.2.1: The Nature of Feasibility Studies

A company will fund a prospecting and exploration if, and only if, they believe there is or will be a market for the commodity that they are seeking. As they begin to invest more time and money into the exploration of the deposit, they will concurrently be evaluating the three categories of factors described in the last lesson. If, at some point, they believe that the project is not worth pursuing, they will most likely stop all work on the project and turn their attention to something else. If at the conclusion of the exploration program, and their “back-of-the-envelope” evaluation, they believe that the project has merit, then they will move to conduct a feasibility study. This is the subject of this lesson.

Practicing engineers and university professors use a variety of adjectives to describe the studies that are conducted, and often do so without complete agreement. Common terms include conceptual (±40%), prefeasibility (±25%), feasibility (±10-20%), or definitive (±5%) studies. The range of terms is meant to convey the cost to perform the study along with the accuracy, (shown parenthetically) of the study. Thus, a conceptual study may cost little to do, but only yield results that are ±40%, whereas the definitive or engineering study in which systems are designed and equipment is selected, will yield highly accurate results. Of course, the latter is laborious and expensive to conduct.

Figure 4.2.1 Cost and Accuracy comparison of different types of feasibility studies

Credit: © Penn State University, is licensed under CC BY-NC-SA 4.0 [1]

The back-of the-envelope considerations described in Lesson 4.1 would qualify as a conceptual or scoping study. I don’t want you to become overly concerned with the words, and accordingly, I suggest you remember it as follows.

The prefeasibility and feasibility studies have a similar goal, which is to estimate the financial merit of the project. Financial merit is quantified by metrics such as return on investment and the number of years until the operation becomes profitable, among other metrics. These studies will be used to make the go/no-go decision on the project. If the money to finance the project is to be raised publicly on the stock exchanges, then the prefeasibility or feasibility study must be made publicly available. The format for the published study is prescribed by law. So, what is the difference between the two studies? Some engineers or companies simply prefer to use one term instead of the other. Perhaps, the meaningful difference between the two terms is the amount of detail.

A certain amount of detail is prescribed by the legal standards for reporting publicly on projects in which investment is being solicited, and that constitutes the minimum level of detail. Often companies will want to invest more time and money into studying the feasibility of the project. The minimum level may be considered a prefeasibility study, and a more detailed examination of the feasibility may be considered a feasibility study. 

Figure 4.2.2 We will think of the analysis as a continuum defined by the degree of certainty or accuracy with three defined regions: conceptual study, feasibility study, and engineering studies. The accuracy of the findings will be on the order of ±40% at the conceptual stage, ±20% at the feasibility point, and ±5 at the engineering point where equipment is being specified and systems are being designed.

Credit: © Penn State University, is licensed under CC BY-NC-SA 4.0 [1]

With this understanding behind us, let’s get on with the subject of this lesson – feasibility studies.

The nature of feasibility studies has been prescribed in by-laws for projects that will seek investments through public listings, e.g., stock exchanges. Most mining projects obtain at least partial funding through public investment, and as such the way we conduct and report on feasibility in a very similar fashion. Let’s take a look at the requirements, which have only come into being over the past few decades.

Since the earliest days of mineral prospecting and development, there has been no shortage of hucksters and shysters attempting to sell worthless mineral deposits. In the old days, the practice of “salting a claim”, i.e., deliberating adding gold or silver to the sample given to the assay office, separated many a person from their hard-earned money. Over the years, the methods to defraud investors became more sophisticated, the size of the investments became orders of magnitude higher, and when fraud occurred, it was likely to affect not just one or two hapless people, but large numbers. In the late 20th century (1990s), there were a couple of huge scams successfully perpetrated, and these caused great losses around the globe to institutional as well as individual investors. Further, it shook the public’s confidence in mining and made it very difficult for mining companies to raise capital.

The major mining countries, including the United States, Canada, and Australia, developed more rigorous standards for public disclosure of scientific and technical information that will be used to solicit investors. In a nutshell, the standards are attempting to ensure:

• the use of common definitions for words like measured resource;
• full disclosure of the people who developed the public report, including names, experience, and potential conflicts of interest; and
• disclosure of information that has a bearing on the worthiness of the project.

There are no significant differences among the standards used by the different countries, although you must ensure that your report meets the standard of the country in which you are going to list your investment opportunity. For example, if you are going to list on the New York Stock Exchange, you must satisfy the U.S. Securities and Exchange Commission’s (SEC) rules, or to list on the Toronto Stock Exchange must comply with the Canadian Securities Administrators (CSA) rules.

We are going to use the Canadian standard because it provides a comprehensive approach that can be easily adapted to the U.S. (SEC), Australian (JORC), or other standards. Even if the intent is not to seek public investment, this standard provides a fine template for reporting on a feasibility study.

4.2.2: The National Instrument 43-101

The Canadian Securities Administrators (CSA) developed and released this legal standard, which is referred to as NI 43-101, in 2001. This standard was built on various codes and policies that had been around for many years, as were the standards developed in Australia and the U.S. The standard provides definitions for qualified persons, feasibility studies, and mineral resource categories. We’ve talked about the last two here in class, but not the first. I’ll leave that reading to you. Please visit this link to the CIM Definition Standards for Mineral Resources and Mineral Reserves [28]. 

As you read through it, take special note of what constitutes a qualified person. Also, note the use of the term modifying factors to describe both the steps to convert a resource to a reserve and the factors that can affect how much of the resource can be converted to a reserve. Finally, you will no doubt be interested in their definition of a pre-feasibility versus a feasibility study. Although it is not part of the standard, the Canadian Institute of Mining (CIM) has prepared and made publicly available a series of best practice documents on the estimation of mineral resources and reserves. Best practice documents are always worth consulting. You may learn something new that you were unaware of, even as a practicing professional. Further, if you deviate from practices that are considered industry-standard practices, you expose yourself to sanctions, lawsuits, and so on. Please visit this Best Practices - Estimation of Mineral Resources and Mineral Reserves [29] document. I suggest that you download it along with the previous document, and save it for future use. At this time, please quickly scan this document to familiarize yourself with the kind of information that it contains so that in the future you will be able to refer to it when you need more detailed guidance.

The “heart and soul” of the standard is known as Form 43-101 F1 [30]. This form essentially specifies the table of contents for the report, and in so doing, is specifying all of the topics that must be addressed in the report. Detailed instructions are given for each of the topical areas. As before, I recommend that you read through this document and save it for future use. 

4.2.3: Examples of NI 43-101 Compliant Reports

Let’s close out this lesson by looking at some examples. These reports have to be made publicly available to investors and their agents, and in many cases, they are available electronically as well as in hard copy. Thus, you can access dozens of these reports freely on the web by doing a search for “NI 43-101”. The quality of these reports will vary. They must meet the minimum requirement of the standard, and some do so – just barely, and are of dubious quality! Some contain a lot more detail than others, i.e., some are definitely preliminary-feasibility studies, whereas others are feasibility studies. There may be times that a company has done more detailed study than it chooses to disclose in the report. That is fine, as long as the additional detail is not considered material. Basically, I am giving you notice that you will see a wide range of detail and quality in these reports that I will give you or that you may find on your own. The list below will give you a nice overview, as it contains different commodities. You’ll want to download and save these reports. You should flip through the pages to become familiar with the type of content; and from time-to-time, you should stop and read a particular section that catches your attention.

National (U.S.) Projects

• Coal: Coal 43-101 elk creek [31]
• Metal: Metal 43-101 gold nugget [32]
• Industrial Minerals:M Potash 43-101 ochoa [33]

International Projects

• Metal: Metal NI 43-101 Rio Tinto Copper [34]
• Coal: Coal NI 43-101 Pitch Black [35]

4.2.4: What's Next?

Let's imagine that we have completed our feasibility study, and the results indicate that this reserve could be made into a profitable mining operation. Given this, we’ll prepare the formal NI 43-101 report. The next step will be to seek financing for our project. If we are going to seek public investment, we’ll list the opportunity on one or more of the stock exchanges. Concurrently, we may be using the report to gain the interest of private equity groups. Regardless, the next step is to obtain financing.

Sometimes the entity that brought the project to this point has no interest in developing and operating a mine. The entity may be a few individuals backed by a venture capital group or may be a company. Regardless, they are interested in turning their hard work into a pile of cash! A large company may want to purchase their reserve to add it to their reserve base – something that they may not mine for several years. Or, a group that does want to work actively to open a new mine may purchase it. For the purposes of our educational journey, we’ll assume that a mining company conducted, or commissioned, the prospecting and exploration and put together the 43-101 report. As such, the company will want to develop and operate the mine.

The engineers on the team will be engaged in time-critical activities while the search for financing is underway. Time is of the essence! A significant amount of money will have been invested to bring the project to this point – perhaps a hundred thousand dollars for a very small project or millions of dollars for a larger one. What have they earned on this investment so far? Nothing, nada, zip! When will they begin to earn something on the investment? Not until they have a mined product to sell. When will that be? Well, for a stone quarry, it may be in as little as a couple of years. In larger projects, eight-to-ten years is not unusual, and in very large and complex projects, it may be closer to two decades! Think about this – sinking money into a project for years and not seeing a penny back!

I have a proposition for you: how about if each of you loans $1000 to me. In return, I will agree to pay you interest on your money, but I will not begin paying anything to you for 10 years. How many of you are going to jump into this investment opportunity that I am offering? How much interest would you need to be promised in order to seriously consider this investment? Oh, and by the way, if things don’t work out quite the way that I plan, I may not be able to pay you back in full, if at all! Welcome to the world of high risk project financing!

Ok, back to the team of engineers doing time-critical things. What are these things that will be happening while financing is being sought? Consider that the time required to bring the project online will include the time that it takes to:

• acquire access rights, as needed
• prepare permit applications
• wait for permit approval
• conduct the detailed engineering studies
• design and bid surface facilities
• construct surface facilities
• order equipment
• provide access to the orebody
• construct underground facilities (if an underground mine)
• hire the workforce

Permits are required by several state and federal agencies. Not only is it expensive and time-consuming to prepare the application packages, but also some of the permits may require the completion of work that may take a year or two. For example, it may be necessary to sample local streams for one year and include the results in the permit application. Once the permits have been submitted, the review and approval process can be tortuous, as the different agencies review and comment on a particular permit. If there is public resistance to the project, public hearings may be required as part of the permit review process. It is important to plan for and sequence the work that you have to do to achieve a timely filing. For example, you don’t want to add an unnecessary delay because you forgot to hire a consultant to conduct an archeological study that is required in support of one of your permit applications!

While the permitting process is underway, detailed engineering studies will be required to design the various systems, e.g., production, materials handling, power, mineral processing, and so on. These designs will then be used to develop specifications for major plant items, such as the mineral processing, equipment and so on. Bid packages will be prepared for surface facilities such as the mineral processing plant, loading facilities, water treatment, warehouses, shops, and so on. If it is an underground mine, the shaft or other access to the orebody will be bid as well. The construction of these facilities can take a few years. So, exactly when do you need each to be completed, and knowing that, when do you need to initiate the bidding for their construction? These are important questions. Given the cost of capital, you don’t want to make the expenditure prematurely; and at the same time, you are going to look rather foolish if mining is set to commence, but the load-out facilities, which are required to get your product off the property and on the way to your customers, have yet to be constructed!

Some equipment can be received within weeks of order, whereas others will be fabricated on-demand, and lead times of several months are common. Choosing the correct point in time to place orders is important.

Accessing the deposit for a surface mine takes less time than for an underground mine, although it can take several weeks to months and must be planned. Underground access can be far more complex. It may take a year to sink a shaft and to develop the spaces around the shaft bottom. And then, depending on the deposit and mining method, it can take weeks to a year or more to develop the workings necessary for mining.

I started this discussion with the question of what are the activities that will be undertaken concurrently with the effort to obtain financing. Clearly, many of the activities that I’ve just outlined will not be initiated until financing is in-hand. However, these activities will be in various stages of planning before financing is completed. Early on, it is crucial that an accurate and detailed project plan be prepared. All of the tasks that need to be performed will be represented on a network diagram known as a PERT diagram. Detailed timing charts, known as Gantt Charts, will be prepared to document the time relationships of the activities that must be completed when they need to be completed and started, and their duration. These diagrams will be used to determine the critical path and when mining can begin, to assess the effect of delays, and ultimately to monitor the progress of the project. Resource requirements will be documented and integrated into the project plan as well. The size and complexity of these projects necessitate the use of project management software, such as MS Project.

By the time financing has been secured, a good project plan will have been completed. Further, work will be well underway to prepare the permit applications; and it is likely that detailed engineering design and analysis will be underway.

The engineering design and analysis of the many systems will be the subject of courses such as MNG 404 Materials handling; MNG 411 Systems Analysis; MNG 422 Ventilation, MNG 431 Rock Mechanics; MNG 410 Underground Mining; and MNG 441 Surface Mining. You will learn more about project management techniques in the capstone design course, MNG 451, where you will conduct a 43-101 feasibility study.

Lesson 4.2 Summary

The next logical step in our progression through this course is to talk about the unit and auxiliary operations. Most mining operations use the same small set of operations to execute the entire mining cycle, and as such, it is convenient to examine them prior to studying each mining method. We will do this in the next Module. Speaking of mining methods, there is one bit of unfinished business before we move on to the next module.

The selection of a mining method is part of the feasibility study. The choice of the method will affect how much of the resource can be recovered, and that, of course, affects the reserve that you can report. When you reviewed Form 43-101F, you saw a section devoted to the mining method, and this is the reason why. Keep in mind that we are not designing the mine at this stage, merely choosing a method based on the data and information we have available to us. In the next lesson, we will look at the set of mining methods and the factors that will influence our selection of a method.

4.3.1: Surface Mining Methods

Surface mining methods are traditionally divided into two classes: mechanical and aqueous. Mechanical methods rely on breaking the ore by mechanical means, and aqueous methods rely on the use of water or another solvent, e.g., an acid, to break down the ore and facilitate its removal.

Mechanical Methods

Open pit mining

This type of mining is used for near-surface deposits, primarily metal and nonmetal. The overburden is hauled away to a waste area and a large pit is excavated into the orebody. The depth of the pit is increased by removing material in successive benches. A few examples of commodities mined by this method would include iron and diamonds.

Figure 4.3.2 Bingham Canyon Mine, November 2017

Credit: J. Sutterlin, © Penn State University, is licensed under CC BY-NC-SA 4.0 [1]

Open cast mining

Open cast mining is also known as strip mining and is used for bedded deposits, and most commonly for coal. Although it is similar to open pit mining, the distinguishing characteristic is that the overburden is not hauled away to waste dumps; but rather, it is immediately cast directly into the adjacent mined-out cut. There are two important sub methods for open cast mining. One is known as area mining, and is applicable when the terrain is relatively flat; and the other is contour mining, better suited for mountains regions. A few examples of commodities mined by this method include coal and phosphate.

Quarrying

Quarrying is a method of extracting dimension stone. The term dimension stone encompasses certain stone products used for architectural purposes such as granite countertops, marble flooring, and monuments, among a few others. The goal in the mining of these products is to remove large slabs that can be cut and machined to exacting architectural applications. Unlike open pit mining in which benching is required to prevent failure of the sides or pit slopes, the high strength and competency of the rock mass in quarries is such that vertical walls of 1000’ or more can be excavated. Now that I’ve given you the classical mining engineering definition of quarrying, you should be aware that just about everybody uses this word, "quarry" to describe any open pit operation in stone! Oh well… A few examples of commodities mined by this method include Georgia marble and Vermont granite.

Auger mining

This is a method to recover additional coal from under the highwall of a contour mine, when the ultimate stripping ratio has been achieved in open cast operations. It is sometimes referred to as secondary mining because it is done after the open cast mine has reached an economic limit.

Aqueous Methods

Hydraulic mining

Hydraulic mining is used for a limited class of deposits that are characterized as loosely consolidated, such as placer-type deposits. A high-pressure water canon is used to dislodge the deposit, and the resulting solution is either pumped to a processing plant or a gravity separation is performed at the mine site using something like a sluice. A few examples of commodities mined by this method include gold and kaolin.

Dredging

This method is used for underwater recovery of loosely consolidated materials using a floating mining machine known as a dredge. In some cases, the deposits are naturally underwater, while in others the area is flooded, creating an artificial lake on which the dredge operates. A few examples of commodities mined by this method include sand and gravel.

Solution mining

Solution mining is used to recover deep deposits that would be uneconomical using underground methods, but only if the ore can be easily dissolved by a solvent. In this method, holes are drilled from the surface into the deposit. A solvent is pumped down one hole, and the resulting solution with the dissolved mineral is pumped out another hole. This solute or pregnant liquor, as it is often known, is processed to extract the mineral of interest. In some cases, only one hole is used, but the hole has an inner and outer section to separate the in-going solvent from the out-coming solute. Water, acid, and steam are common solvents. A few examples of commodities mined by this method include uranium and sulfur.

Heap leaching

Heap leaching was used many years ago as a method to recover very low percentages of metal remaining in the tailings from mineral processing plants. Large piles, i.e., heaps, of the tailings of low-grade ore were created, a solvent was allowed to drip and percolate down through the heap, and then the pregnant liquor was recovered and processed. In this fashion, it is a secondary method. In recent years, it has been used with increasing frequency to recover high-value metals such as gold from very low-grade ores. A few examples of commodities mined by this method include copper and gold.

4.3.2: Underground Mining Methods

Underground mining methods become necessary when the stripping ratio becomes uneconomical, or occasionally when the surface use of the land would prohibit surface mining. Underground methods are traditionally broken into three classes: unsupported, supported, and caving methods. These classes reflect the competency of the orebody and host rock more than anything else. If you excavate an underground opening in the ore or the rock, is the opening stable -- i.e., will it remain open for an extended period, or will it begin to fall in? If it is unstable, i.e., the surrounding ore or rock breaks up and falls into the opening, how much support would be required to keep the opening from caving in? The answers to these questions lead us to choose mining methods from one of the three classes. Unsupported methods require the addition of minimal artificial supports to secure a stable opening, whereas the supported methods require the addition of major support to keep the openings from caving in. Finally, the third class is, at first glance, counterintuitive: in general our goal is to create stable openings underground for obvious reasons, but the methods in this class will only work if the host rock or orebody will cave easily under its own weight -- the caving methods actually depend on this caving action to function safely and productively!

Unsupported Methods

Room and Pillar mining

This method of mining is used to recover bedded deposits that are horizontal or nearly horizontal when the orebody and the surrounding rock are reasonably competent. Parallel openings are mined in the ore, i.e., rooms, and blocks of ore, i.e., pillars, are left in place to support the overlying strata. Other than the pillars, little artificial support is required and often consists of bolts placed into the overlying strata to pin the layers together, making them behave like a strong laminated beam. A few examples of commodities mined by this method would include coal, lead, limestone, and salt. Historically, if the pillars were irregular in size and placement, which is more likely to occur in certain metal and nonmetal deposits, this method was known as stope and pillar, rather than room and pillar. You will still hear the word stope and pillar being used, but the distinction is now largely irrelevant. This method accounts for the vast majority of all underground mining in the U.S., and likely the world.  Watch this video (2:58) created by Caterpillar showing the use of their equipment in room and pillar mining. 

Shrinkage stoping

Shrinkage stoping is used to recover steeply dipping orebodies when the ore and host rock are reasonably competent. A stope, i.e., a large section of the mine where active production is occurring, is mined, but the broken ore is not removed, but rather is left in place to support the walls of the stope until the time when all of the broken ore will be removed. Since rock swells, i.e., increases in volume when it is broken, it is necessary to draw off some of the broken ore as the stope is progressively mined. The name of this method derives from this drawing off or shrinkage of the stope. A modern and important variant of this method is known as vertical crater retreat (VCR) mining. A few examples of commodities mined by this method include iron and palladium. Watch this video (3:01) created by Atlas Copco demonstrating sublevel stoping mining method.

Open stoping

This type of mining is used to recover steeply dipping orebodies in competent rock. The ore is removed from the stope as soon as it is mined. Sublevel stoping and big-hole stoping are the important variants in use today. A few examples of commodities mined by this method include iron and lead/zinc.

Supported Methods

Supported methods historically included cut and fill stoping, stull stoping, and square set stoping. However, the last two are no longer used due to their extreme cost. We’ll confine our discussion to cut and fill stoping.

Cut and fill

Cut and fill is used to recover ore from weaker strength materials, in which the openings will not remain stable after the ore is removed, and the overlying strata cannot be allowed to cave. A slice of the orebody is mined and immediately after the ore is removed, backfill is placed into the opening to support the ore above. The next slice is removed, the cut is then backfilled, and the process repeats. As you might imagine, this is a very expensive method to use, and consequently, it would be used only for the recovery of high value ores. An example of a commodity mined by this method is gold. Watch this video (2:58) created by Altas Copco on Cut and Fill mining method.

Caving methods

Caving methods include block caving, sublevel caving, and longwall mining. For emphasis, allow me to repeat what I said earlier: caving methods are used in settings where the ore or the host rock is so weak that it cannot support its own weight for any period of time; the methods only work if the rock or the ore will readily cave under its own weight.

Block caving

This method is used in weak and massive orebodies, in which the ore is undercut, and then as the broken ore is removed the remainder of the orebody collapses into this void, and as more ore is withdrawn, the caving continues. Typically the host rock is fairly strong, although ultimately it tends to cave into the void created from removing the ore. The fracturing and caving often break through to the surface. Watch this video (3:16) created by Atlas Copco on Block Caving Mining Method

Sublevel caving

This type of caving is used in strong and massive orebodies in which the host rock is very weak and quickly caves into the void created by removing the core. As in block caving, the cave will ultimately reach the surface. Watch this video (3:05) created by Atlas Copco on sublevel caving mining methods.

Longwall mining

Longwall mining is a type of caving, applied to a horizontal tabular deposit such as coal. While block and sublevel caving are essentially vertically advancing metal mining methods, longwall mining is applied to relatively thin and flat-lying deposits – most often coal, but occasionally an industrial mineral such as trona. The coal seam is extracted completely between the access roads, and then as mining retreats, the overlying strata caves into the void left by removing the coal.  Watch this video (5:31) created by Clearcut Mining Solutions showing logwall mining method.

Our goal in attempting to classify mining methods is to make it easier to learn the methods, because methods in a given class tend to work best in similar circumstances. Similarly, there tend to be just a few factors that differentiate the methods. By examining the classification scheme, we make it easier to remember the methods and the characteristics under which they can or cannot be used. It’s also useful to note that there is nothing sacred about the choice of a method. If five years down the road the characteristics of the deposit are changing, then another method will be employed. There are examples of mines utilizing three different mining methods over a 15-year period, as they adapt the mining method to the evolving geological conditions. Sometimes, one method is employed as the primary mining method, but another is used on retreat to recover pillars, for example. We’ll look at some of those cases later as well.

4.3.3: Why Do We Choose One Method over Another?

There are many factors that can affect the choice of a mining method. However, a relatively small number of them will dictate the choice. The others may affect the layout of that method, or other details, but rarely do they eliminate a method from consideration or drive the selection of a method. Let’s take a look at a comprehensive set of factors and understand what they mean. Then, we’ll step back, take a deep breath, and see how uncomplicated it can really be! Here is a comprehensive list with a few annotations to indicate the significance of the factor.

Spatial characteristics of the deposit

These factors play a dominant role in the choice of a mining method because they largely decide the choice between surface and underground mining, affect the production rate, and determine the method of materials handling and the layout of the mine in the ore body.

1. size (especially height, thickness, and overall dimensions)
2. shape (tabular, lenticular, massive, or irregular)
3. attitude (inclination or dip)
4. depth (mean and extreme values, stripping ratio)
5. regularity of the ore boundaries
6. existence of previous mining

Geologic and hydrologic conditions

Geologic characteristics of the ore and surrounding country rock influence method selection, especially choices between selective and nonselective methods, and ground support requirements for underground mines. Hydrology affects drainage and pumping requirements, both surface and underground. Mineralogy governs solution mining, mineral processing, and smelting requirements.

1. mineralogy and petrography (e.g., sulfides vs. oxides in copper)
2. chemical composition (primary and secondary minerals)
3. deposit structure (folds, faults, discontinuities, intrusions)
4. planes of weakness (joints, fractures, shear zones, cleavage in minerals, cleat in coal)
5. uniformity of grade
6. alteration and weathered zones
7. existence of strata gases

Geotechnical (soil and rock mechanics) properties

The mechanical properties of ore and waste are key factors in selecting the equipment in a surface mine and selecting the class of methods (unsupported, supported, and caving) if underground.

1. elastic properties (strength, modulus of elasticity, Poisson's ratio, etc.)
2. plastic or viscoelastic behavior (flow, creep)
3. state of stress (pre-mining, post-mining)
4. rock mass rating (overall ability of openings to stand unsupported or with support)
5. other physical properties affecting competence (specific gravity, voids, porosity, permeability, moisture content, etc.)

Economic considerations

Ultimately, economics determines whether a mining method should be chosen, because economic factors affect output, investment, cash flow, payback period, and profit.

1. reserves (tonnage and grade)
2. production rate (output per unit time)
3. mine life (total operating period for development and exploitation)
4. productivity (tons /employee hour)
5. comparative mining costs of suitable methods
6. comparative capital costs of suitable methods

Technological factors

The best match between the natural conditions and the mining method is sought. Specific methods may be excluded because of their adverse effects on subsequent operations (e.g., processing, smelting, environmental problems, etc.).

1. recovery (proportion of the ore that is extracted)
2. dilution (amount of waste that must be produced with the ore)
3. flexibility of the method to changing conditions
4. selectivity of the method (ability to extract ore and leave waste)
5. concentration or dispersion of workings
6. ability to mechanize and automate
7. capital and labor intensities

Safety, Health, Environmental concerns

The physical, social, political, and economic climate must be considered and will, on occasion, require that a mining method be rejected because of these concerns. The impact of one mining method over another method on the environment must be considered. Similarly, the ability to provide the highest level of safety and health with one method as compared to a competing method must be considered.

1. ground control to maintain integrity of openings
2. subsidence, or caving effects at the surface; or affect on groundwater system
3. atmospheric control (ventilation, air quality control, heat and humidity control)
4. availability of suitable waste disposal areas
5. workforce (availability, training, living, community conditions)
6. comparative safety conditions of the suitable mining methods

4.3.4: Putting it into Perspective

So there you have it – the 37 factors that will influence your choice of a mining method… but how and when? Fortunately, this all reduces to a few major drivers.

First is depth...

If I know the depth of the deposit and the thickness of the overburden, I can do a few calculations and decide whether it is most likely going to be a surface or an underground mine. With this one factor, I’ve excluded or included half of the mining methods. Here, our decision tree has to split based on surface or underground. Let’s go down the surface path first.

If it’s a near-surface deposit, then tell me if it's metal, nonmetal, or coal deposit. If it’s a noncoal deposit, then open pit is likely. If it’s coal, then open cast is likely.

If it’s coal, then tell me about the topography. If it is flat lying, area mining is likely. If it is mountainous, then contour mining is the better choice.

On the other hand, if it is a low-grade and deep deposit, then solution mining will be considered if the mineral is one that is known to be recoverable with solution mining methods.

If the deposit is dimension stone, then I know it is going to be a quarry operation.

Ok, so what if an underground method is indicated?

The process is not quite as simple as for narrowing the field of surface methods, but almost so. Let’s go down that path and see.

First, I’d like to know about the attitude. Is the deposit horizontal or nearly so? If so, I’ve excluded several of the underground methods, e.g., shrinkage stoping and open stoping. On the other hand, if it is steeply pitching, I can eliminate room and pillar.

Next, I’d like to know about the competency of the host rock and the deposit. That will further narrow the field of potential methods.

Figure 4.3.3 Sample Decision Tree for determining the Mining Method after depth of deposit and thickness of overburden is calucluated

Credit: © Penn State University, is licensed under CC BY-NC-SA 4.0 [1]

This will all become clearer...

After we’ve studied the methods in more detail, this will become clearer. At this time, I am simply trying to make the process of selecting a method seem less intimidating. Sure, all of the 37 factors that I listed earlier are relevant, and that will become apparent by the end of the course. The ones with the greatest effect, in general, are:

• depth;
• surface features (especially for, but not limited to surface methods);
• attitude, shape, and size (extent);
• geotechnical characteristics of the orebody and the host rock;
• grade and uniformity;
• production requirements.

If these are the factors that essentially drive the selection process, then why do we bother listing the others? You will be in a stronger position to answer this at the end of the semester, but let me make a few remarks now, to give you a better feel for the relevance of the other factors, and why you should learn them!

• Safety considerations, as well as productivity, are responsible for the preference of vertical crater retreat mining rather than shrinkage stoping.
• Selectivity will be important for high grade ores in which the precious metal is concentrated in narrow veins or bands.
• Presence of discontinuities could nix the use of one caving method, but be largely irrelevant for a different caving method.

Lesson 4.3 Summary

In this lesson, we’ve introduced the different mining methods used to exploit mineral deposits. We’ve characterized them into broad categories of underground and surface. Within each category, we established classes of methods, and then we identified the individual mining methods belonging to each class. We saw that a class represented a few specific characteristics of the deposit, and as such the methods in that class are well suited for deposits with those characteristics.

The choice of a specific mining method may require consideration of several factors, and we looked at six groups of factors totaling 37 in all. Although any of these factors can affect the selection of a mining method, a small set of the 37 have a disproportionate effect on the choice, and we identified those.

You will develop a better understanding of the details of the mining methods and the many factors that affect the choice of the method as we work through this course. Despite the lack of detail at this stage, you will find the material covered in this lesson to be quite useful as we continue into the remaining modules.

Regardless of which method we use, it is likely that the similar unit and auxiliary operations will be used during exploitation. The equipment itself may be very different, but the operations are similar from an engineering perspective. We’ll take a look at this in the next module.

5.1.1: Equipment for Winning the Ore

Commonly used equipment for winning the ore includes the items listed below. The selection of a specific size or type of equipment generally requires an engineering analysis. Here, the goal is simply to familiarize you with the name and function of the major pieces of equipment.

5.1.2: Materials Handling

Materials handling is concerned with how we move (haul) the mined material out of the mine to the processing plant. Before looking at specific haulage options, we should talk about a few overarching concepts. The first is intermediate haulage.

At the working face, i.e., where the ore is being freed from the deposit, we are going to load the ore into some type of haulage. If we are in a surface mine, it is likely that we’ll load the ore into a truck, which will haul the ore the entire way to the plant, dump it, and then return to the face for another load. If we are in an underground mine, it is likely that the material at the face will be loaded into a haulage vehicle and then transported to an intermediate dumping point. From the intermediate point, a different haulage device will be employed. Sometimes, there could be even another subsequent intermediate device. As an example, assume we have a mine in which LHDs are used to load out the ore at the draw point (face) and transport it to a dump. The dump point is at an ore chute in the rock that funnels the ore to a lower level where it is loaded into rail cars, along with the ore from many other dump points. A train may pull these cars several miles out of the mine and to a processing plant. Or, the train may take the ore to a transfer point at the bottom of the shaft, where the ore will be dumped and transported in large ore skips (buckets) up the shaft to the surface. The choice of a specific type of materials handling is based on optimizing the overall process. Smaller and maneuverable equipment is best suited at the face, whereas larger capacity, but more permanent equipment may be indicated to move the material out of the mine. As we look at different mining methods, this will become even clearer.

Let’s identify the common choices for materials handling, to move material from the face to the plant.

Figure 5.1.21 Materials handling equipment, shown in three groups
Click here to see a text description.

• Group 1
  • LHDs
  • shuttle cars
  • bridge conveyors
• Group 2
  • mine trucks
  • haul trucks
  • rail
• Group 3
  • rail
  • haul trucks
  • hoists
  • hydraulic transport
  • belt conveyors

Credit: © Penn State University, is licensed under CC BY-NC-SA 4.0 [1]

I’ve arranged them into three groups. The equipment in the first group is typically used for short haulage runs from the face to an intermediate dump or transfer point. The equipment in the third group is used to move the material out of the mine and to the plant. The equipment in the second group can fall into Groups I or III. In some instances, the equipment shown in Group II will be used to move the material directly from the face to the plant, whereas in others, it will be used to transfer the ore to an intermediate point. Again, the choice and rationale will become clearer as we learn about the requirements inherent to the different mining methods. Let’s look at the general characteristics of these materials-handling modes. We’ve already discussed LHDs, so let’s begin with shuttle cars.

Lesson 5.1 Summary

There is, of course, much more that we could say about materials handling systems. Indeed, we have an entire course on their design! For our purposes here, this brief overview will help us to understand the sequence of operations in the mining methods that we will be studying.

Knowledge Check 

In Lesson 5.3, we’ll take a look at the difference between batch and continuous operations, and the evolution of the so-called “continuous mining systems.” However, before talking about that topic, we will look at auxiliary operations next.

5.2.1: Ground Control 

An Overview of the Requirement

One of the more important and defining characteristics of mining engineers is their ability to design structures in earth materials. Few appreciate the incredible challenge this presents. Consider that if we want to design a building, we can go to a handbook and find the exact mechanical properties of our building materials, whether they are steel beams, lumber, or concrete. Important engineering challenges may remain in the design of the building, but at least we know the properties of our building materials.

…not so when we are constructing something in the crust of the earth. Sure, we can conduct lab tests and that is helpful, but significant uncertainty remains throughout the property in many cases. Even when we have a well-defined material, a limestone for example, it may have structural defects due to geologic processes that result in fractures, joints, and so on. Once past the hurdle of knowing the properties that we are working with, we have to work with them! That is to say, some of the materials may be “inferior” or “substandard” to use construction terminology. But, in our case, we can return those “inferior” rock structures to the maker, and demand new ones with better properties. And here lies what can be a daily challenge for the mining engineer: maintaining a stable and safe rock mass to allow mining of the valuable commodity.

Whether we are designing and operating a surface or underground mine, this will be a challenge of varying degrees. In the surface mine, we are concerned with a failure of the slope or the highwall. If it is too high or too steep or if there has been an influx of groundwater, or… we can have a failure. The failure could kill all of the workers in the area, destroy or bury the equipment, and result in costly production delays. In an underground mine, we develop a series of openings to access and exploit the deposit. If any of those openings cave in, it can have disastrous consequences for the miners and the overall viability of the operation. The art and science of safely designing structures in the earth is captured in the term ground control. The coursework in engineering mechanics, strength of materials, and rock mechanics will give you a solid theoretical foundation, and then courses in ground control along with years of experience will enable you to succeed in this endeavor.

I think you are beginning to appreciate the importance of this auxiliary operation, which is an intimate companion with the unit operations of the production cycle. Here in this course, we’ll talk about how the ground conditions affect the mining methods and the sequence of operations, and some of the practices that are used to “control the ground” to prevent unwanted and unexpected failures.

Ground control is achieved in part through the design of the structures. For example, the engineer will design spans that won’t fail or determine the angle of the highwall to reduce the risk of a slope failure. We’ll leave the design component of ground control to other courses. Operationally, after the design, ground control is achieved in part through the use of various technologies. Uses of these technologies are an ongoing part of the production cycle, and as such are of special interest in this lesson.

Cyclical Practices

After drilling and blasting, the integrity of the surrounding ground will be inspected. This will happen prior to loading to ensure that miners engaged in the loading operation are not put in danger. This check is imperative in underground mining, and in many surface mining operations. There may be pieces of rock that have cracked but hung up, and they may fail without warning. Or there may be loose pieces that could fall. Thus, the inspection will be the prelude to scaling or barring, i.e., the act of knocking these pieces free. In a low seam mine, with minimal need for scaling, this may be accomplished with miners using a steel bar with a chiseled edge. When it is impossible to reach the area that requires scaling, then equipment will be used. Here is a scaler that is found in metal/nonmetal mines; and primarily but not exclusively in underground mines. It has a telescoping boom to reach high places and a pneumatic hammer to release any worrisome pieces of rock. The operator remains inside the cab, which provides protection from falling material. Notice also the dozer blade on the front. It is used to push a fallen material out of the way.

Ok, so this inspection and scaling represents the first auxiliary operation.

Figure 5.2.1: A scaler

source: Thiessen Team Rock Solid Solutions [51]

Once the area has been scaled, it will be safe to bring in equipment for loading and hauling in some mines – mines with competent rock that is not going to cave. In some mines, it is necessary to take further steps to improve the opening to ensure that the rock strata is not going to fall onto the miners, and this must be done before miners work in the freshly mined space. If additional reinforcement of the rock is required, then that will be the second auxiliary operation.

We’ll take a look at some of the equipment used to reinforce the rock, but first, a quick tutorial on the options for adding additional support to the ground.

5.2.2: Active and Passive Ground Support

Ground support can be active or passive. If we take actions that actually increase the strength of the rock, then we are providing active support. On the other hand, if we do nothing to increase the strength, but instead we take an action to prevent the rock mass from falling, then we are providing passive support. There are valid applications of both categories of support. Let’s look at an example. Actually, this is a really important example!

The rock strata overlying a tabular deposit are often somewhat weak. They may consist of relatively thin layers of a weak rock. These layers, which may be a few inches to a few feet in thickness, may also have partings between them. These partings may be very thin, and very weak. Now, what is going to happen when we remove the ore from underneath the overlying rock mass that I just described?

Did you ever have a long bookshelf made of wood – wood that is not very thick? What happened when you place a full load of books on the shelf? It begins to sag in the middle, right? It may not have happened immediately, but over time, it sags in the middle. A similar mechanism is at play in the mine strata of our example. All of the weight of the overlying rock layers is pushing down. You removed the ore from underneath the strata, so you now have a long unsupported span of relatively weak rock layers. The rock layers begin to sag, and before long the lowest layer will separate from the others, and fail. When it fails, it will break apart and fall to the ground. Now, the layer immediately above the one that just failed will sag, and it too will fail. This will continue until the rock mass reaches some equilibrium. This is an unacceptable outcome! We’ll never make any money at this mining business if we can’t support the openings that we create! So, you must prevent the initial failure. What are your options?

As with the bookshelf, you could place a support in the middle of the span to support the load and prevent sagging. In our mine, we could set a timber post into place, and it would prevent the failure. This is an example of a passive support. Placing timber posts, steel arches, and so on is a time-honored means of supporting the rock and preventing failures. There are some drawbacks, however. If we’re mining a 6’ thick coal seam, we can set a 6’ post. What if we are mining a 60’ seam of limestone? Finding, handling, and placing a 60’post would be a supremely daunting task, and clearly not practical! Even when it would be easy to set a post, there can be unacceptable drawbacks to the practice.

Let’s think about a basketball arena like Bryce Jordan Center. The roof of that building spans a large area, and that roof needs to be supported. Assuming that we can obtain timber posts of sufficient length, why not set them across the width of the building, perhaps every 20’ or so to support the roof? For starters, it would make for an interesting new twist to the game of basketball as the players dodged the posts on their way to the basket! Again, not very practical, and that is the drawback in the mine as well. Instead of basketball players, we have large equipment moving around and between the working areas. Thus, in many locations, this kind of passive support is inconsistent with production needs. While passive support is an appropriate control strategy in many instances, we need an alternative for when it is not.

Roof Bolts

One of the most prevalent and effective means of providing ground support is an active support known as a roof or rock bolt. Remember the failure in the layered strata of our example mine? What if we were to clamp several of those weak layers together, forming a “laminated beam”? Beams are very strong, and, indeed, by clamping these layers together, we have strengthened the rock mass, and it will be able to support the weight on it without failing. Many of you may have seen laminated beams used in building construction. Several layers of thin and relatively weak plywood are glued together, and the result is beam with superior properties to a solid piece of wood of the same thickness.

In our case, we can’t access both sides of the rock mass to clamp the layers together, as is the case when they fabricate laminated wooden beams. Instead, we use a rock bolt, as illustrated in this diagram.

Figure 5.2.2: Diagram of a roof bolt

Credit: © Penn State University, is licensed under CC BY-NC-SA 4.0 [1]

It works as follows. A hole is drilled into the rock mass that is to be supported. Next, the roof bolt is inserted into the hole. Then the bolt head is rotated, which causes the expansion shell to expand into the surrounding rock, locking it in place, and as the bolt rotates, it draws in the bearing plate. As a result, the layers between the expansion shell and the bearing plate are compressed tightly together, forming a beam. For this to work, the expansion shell must be anchored in competent rock. Sometimes, this can be done over 4’ and other times, 6’ or more. The length of the bolt is chosen with this in mind. The diameter of the bolt, which is between ½ - 1”, is based on the required tensile strength of the bolt, which is determined by calculating the weight of overlying rock that each bolt will have to support. In the future, you will learn how to size rock bolts. At this time, the takeaway message is:

1. bolting is an active support because it is increasing the strength of the rock mass to prevent failure by clamping weak layers together to form a strong beam; and
2. bolting is the most prevalent means of ground control in the majority of underground mines, and also in certain surface mines.

Under difficult conditions, we can place a skinny bag of resin, i.e., a strong glue, into the hole with the bolt; and then, as the bolt spins while it is being tightened, the bag will break and the resin will be mixed and will completely fill the hole around the bolt. Moreover, this resin will plug cracks and fractures in the rock in the immediate vicinity to further enhance the bolt’s holding power.

Here is a picture of roof bolting machine, commonly known as roof bolter or simply, a bolter. If you look closely, you will see the drill in front of the operator. Notice the attachment point to the machine at the bottom, i.e., the chuck. As the hole is drilled, that chuck assembly will move up towards the collar of the hole. After the hole is drilled, the operator will place a bolt in the chuck, insert it into the hole, and tighten it. The machine has plenty of space to store a supply of bolts along with the drill tools. You will also notice hydraulic canopies pressing into the top or “roof” of the mine opening. Those are passive supports to ensure that the top does not fall down while the bolter operator is installing the roof bolts.

Figure 5.2.3: Roof bolting machine

JH Fletcher & Co. [52]

In so-called large opening mines, where the distance between the bottom and top of the mine opening might be 20 – 30’ or more, a slightly different configuration is required, as shown here. The operator would stand in the “cherry picker basket" and operate the drill and bolter.

Figure 5.2.4: Roof bolter for large opening mines

JH Fletcher & Co. [53]

Cable Bolts

Sometimes, we need a bolt that is much longer than 4 – 6’. Remember, I said that for the bolt to achieve its purpose, the anchor must be in a competent rock layer. What if that layer is 15’, or 50’ or even 100’ above the opening? A bolt of such lengths would be impractical. Instead, we use a cable bolt. A reel of specially constructed steel cable is stored on a cable-bolting machine. A hole of the required length is drilled, an anchor shell is attached to the end of the cable, and it is fed to the end of the hole. An anchor plate and tensioning bolt is affixed to the end, and it is tensioned in some fashion as a traditional rock bolt. One difference, however, is that concrete grout will be pumped into the hole as well. Here is an example of a cable bolter. If you look closely, you can see the cable routed from the spool on the rear of the machine, up over the machine, and to the front where it will be inserted into the hole. The carriage housing the drill steel and the drill are also visible at the front of the machine.

Figure 5.2.5: Cable bolter

Enter image credit here

Shotcrete

Shotcrete is another active support of note. Pillars of rock or ore are often left standing to help support the overlying strata, and only the ore between the pillars is recovered. The pillars can be under a lot of stress, and in some cases will begin to fail. The outer layers tend to crack and spall off, until the pillar is much skinnier and more likely to fail. The spalling of the outer layers can also present an immediate safety problem. One means of active support is to spray a Shotcrete onto the pillar. The Shotcrete may be cement or polymer mix, and fibers of various materials may be embedded in the mix. Thicknesses of an inch or more would be typical. Once the Shotcrete has dried, or really cured, it is quite strong and provides a confinement stress on the pillar. Sometimes the top surface of the span will be shotcreted as well.

Here is a picture of a shotcrete operation in progress. You may be able to see the nozzle at the end of the boom, the mixing tank, and the dry feed hopper, and water line on the equipment. The other line is an electric power cable. The hose and cable reels provide sufficient mobility to the machine to allow it to move freely among working places, at least until the limit of the hose and cable are reached. Then, it will be necessary to advance the electrical and water supplies. Advancing utilities, such as electric power and water, is another important auxiliary operation that we will discuss shortly.

Figure 5.2.6: Concrete sprayer

AtlasCopco.com [54]

We’ve covered the majority of the ground control technologies with this brief introduction to active supports such as roof bolts, cable bolts, and shotcrete; and passive supports such as timber or steel posts and arches. There are others, but this group is illustrative. There is one more ground control technique that deserves a brief mention before we move on to other auxiliary operations, and that is backfilling.

Backfilling

Backfilling is the process of adding support to a mined out area by filling it with backfill, i.e., waste rock, tailings, or a mixture of cement and tailings or waste rock. As you might suspect, the time and cost to backfill is very expensive. Nonetheless, it is a part of the cut and fill mining method, in which we routinely mine a slice of ore, backfill it, and mine out the next slice, and so on. We’ll talk more about backfilling when we discuss this mining method. For now, you should simply be aware of backfilling as another ground control practice.

Also, I want to be clear that while the ground control challenge in underground mining is generally far more complicated than in surface mining, ground control practices including inspection and scaling are part of surface mining, and the use of rock bolts is part of some surface operations.

After this brief tutorial of ground control, you’ve probably forgotten how we got started on this discussion in the first place! Ground control is a critical auxiliary operation, and although we usually speak generally of ground control, in total, we’ve now seen that there are three components: inspection, scaling and installing ground support. Depending on the mining method and the competency of the rock, the activities associated with ground control may occur at differing time points in the production cycle.

5.2.3: Electrical Systems

The power sources for mining equipment are compressed air, diesel, and electric. Most mining equipment today is powered with either diesel or electric. Diesel provides more mobility because there is no need to be tethered to an electrical cable. On the other hand, the exhaust from a diesel engine creates challenges when used in a confined space, like an underground mine, and it requires a high level of regular and sophisticated maintenance. Electric has certain inherent advantages including the ability to provide a lot of power in a small space and no toxic emissions at the machine. Large equipment, such as draglines and shovels, as well as continuous miners, is usually electric. Diesel is favored for haul trucks because of the need to be untethered.

Nearly fifty years ago, there was an effort to replace electrically powered equipment with diesel engines. Eliminating the delays associated with the tether, as well as delays associated with inevitable failures of the tethers (cables) was easy to justify; and particularly so for mobile materials handling equipment. In recent years, the air quality issues associated with diesel particulate matter and noxious gases from the engine exhaust have caused some operators to transition back to electric equipment.

I should mention battery-powered equipment, given that today we are seeing more and more battery-powered vehicles on the highways. With battery power, the tether, i.e., the cable, is eliminated, and there aren’t the drawbacks associated with diesel engines. Unfortunately, it is difficult to pack sufficient energy into a battery pack to replace the “workhorses” of hauling ore and rock. Consequently, battery-powered equipment is used for utility vehicles, such as personnel carriers or scoops for hauling supplies.

The machinery associated with production will move throughout the mine over the mine life, and as this equipment makes its way through the deposit, it will be necessary to move the electrical power system along with it. Depending on the mining method, we may need to move power every few shifts, or in other cases, only every few months. Many of you have some familiarity with power systems found in towns – the substations, power lines, and so on. Can you imagine having to move and advance such a power system on a frequent basis? In many mining methods, that is all part of a day’s work! Of course, we’ve taken steps to make it more manageable. The substations and other equipment that have to be moved are mounted on skids, crawlers, or rubber tires to facilitate movement. Couplers are used to connect and disconnect components and cables. It’s manageable!

Of course, there are other electrical loads, some of which must be moved periodically, and others that are stationary and semi-permanent. These loads would include pumps, conveyors, and illumination. Even if the mining machinery is all diesel powered, it is likely that there would be other electrical loads that must be powered. Therefore, providing electrical power will be a necessary auxiliary operation for most mining methods used in surface and underground mining.

Here is a picture of a skid-mounted power center for use in an underground mine. You can see the high voltage cable going into the box. Inside, there is a transformer to reduce the distribution voltage, typically 4.16 to 15 kV, to the level required by the machinery, typically 480 to 995 V. On the other side of this box, outside of our view, there are couplers to connect the various loads. Protective devices for each outgoing circuit are built into the power center.

Figure 5.2.7: Skid mounted power center

Intermountainelectronics.com [55]

Here is a picture of a similar power center, which has just been built. Here, you can get a better view of each outgoing circuit. Each cable-connected load plugs into one of these circuits. The lid of the power center is removed only for maintenance purposes.

Figure 5.2.8: Newly built power center — better view of the outgoing circuits.

Source: PEMCO, used with permission.

You might see a load center like the one in the following picture in a surface mine. Essentially, a shipping container has been mounted on a skid, and the required electrical components have been assembled inside the container. I’ve seen power centers in surface mines that have been built on railroad cars and old tractor-trailer beds, among others!

Figure 5.2.9: Surface mine load center

Meglab [56]

As I mentioned earlier, these power centers must be advanced as mining progresses. Often a “power move” will require adding additional high-voltage cable, or in some surface mines, adding additional poles and overhead lines. Some mining methods are more “electric-intensive” than others, and that also varies by commodity. Room and pillar mines in coal, salt, or trona, for example, are very power intensive, requiring significant infrastructure and frequent moves. Others, such as room and pillar mining in limestone or cut and fill mining in gold, require less electrical infrastructure and few moves. Nonetheless, providing safe, reliable, and timely power is a critical auxiliary operation in most mines.

5.2.4: Ventilation

Fresh air is required to provide oxygen and to carry away carbon dioxide. Large quantities of air are often required to dilute, render harmless, and carry-away dangerous dusts and gases. Methane is an explosive gas associated with coal deposits, among others. Sufficient air must be coursed through the coal mine to keep the methane concentration well below the lower explosive limit. Hydrogen sulfide, a deadly gas, forms under certain conditions in underground metal mines and must be diluted to render it harmless. Carbon monoxide and oxides of nitrogen are produced during blasting, and these toxic gases must be diluted and carried away before miners return to continue the production cycle. The concentration of respirable silica dust, i.e., dust small enough to become trapped in the lungs, is produced during mining and can lead to fatal lung diseases when inhaled over years of work. Accordingly, the allowable concentration of respirable dusts is heavily regulated and controlled, and providing sufficient air to carry away these dusts is important. We could go on with the examples, but you get the idea!

It is easy to understand the need to provide large quantities of air, sometimes on the order of several hundred thousand cubic feet per minute! It is not so easy to control that air, to get the right amounts to the different parts of the mine, and then to move the “dirty air” to outside of the mine. It takes significant engineering, and many control devices, to achieve the intended outcome. If you look, for example, at this plan view of a mine, below, you can see the many passageways. It resembles the streets and avenues of a city. Now imagine that you have one or two entry points for fresh air, as provided by fans, and you have a few exit points where the exhaust air is ventilated to the outside. All that you have to do is to make sure that the proper quantities are flowing in each of the “streets” and “avenues.” Further, you have to do this on an ongoing basis, since the layout of the mine changes from day-to-day as mining advances. No small feat…! Depending on the mining plan, you can build stoppings, i.e., solid barriers to divert air, you can hang curtains to divert air, you can install and adjust regulators to control pressure drops and the attendant airflows, and you move auxiliary fans and tubing to further direct airflows to where they are needed.

When we looked at ground control, we identified the importance of inspection, as the first step in that auxiliary operation. Similarly, with ventilation, there is a “check” that occurs as a prerequisite to certain unit operations. Gas measurements may be required as well as air quantity determinations. It may be necessary to adjust the ventilation system as a result of these measurements or to change the production plans until adequate ventilation can be provided.

Our discussion has focused on ventilating underground mines, and indeed that’s where most of the action is to be found. Interestingly, surface mines occasionally experience challenging ventilation problems. Some open pit mines, for example, are deep enough that temperature inversions can occur, in which air in the pit is trapped by a cooler and heavy layer of air near the top of the pit. The diesel fumes from the heavy equipment operating down in the lower levels of the pit can build up to dangerous levels. Utilizing artificial ventilation in this circumstance, whether the wind induced from the blades of a helicopter hovering over the pit or large axial vane fans, is essential to protecting the health of the miners and ensuring that production can continue.

Ventilation, as with ground control, is a crucial auxiliary operation.

Figure 5.2.10: Ventilation system

Credit: J. Kohler, © Penn State University, is licensed under CC BY-NC-SA 4.0 [1]

5.2.5: Maintenance

The level of maintenance required to keep the equipment operating properly and safely varies by the mining method and the particular equipment in use. Usually, the idea is to minimize breakdowns during a production shift and the need for emergency maintenance. Instead, preventive maintenance is practiced to reduce failures, and scheduled maintenance is planned for nonproduction shifts. In some mines, there will be two eight-hour production shifts followed by an eight-hour maintenance shift every day. In other mines, there will be two ten-hour production shifts, followed by a four-hour maintenance shift. And in yet other mines, a week of production will be followed with a weekend of maintenance.

The value of a minute is often underestimated by those just learning about the industry. Given the capital cost of mining equipment, and the cost of labor and so forth, a delay can easily cost hundreds of dollars per minute and exceed a thousand dollars a minute in certain circumstances. As such, maintenance is a critical auxiliary operation.

There is an important metric that is used to keep track of how well the mine is doing with its maintenance. The metric is known as availability, which is defined as the amount of time that the equipment was usable divided by the amount of time that we had planned to use it. For example, suppose that we planned to install roof bolts for six hours yesterday. A hydraulic hose ruptured, and it took 30 minutes to obtain and replace the hose. The availability of the roof bolter yesterday is: $$\left(\frac{360\min -30\min }{360\min }\right)=9.166\approx 92\%$$ Instinctively you may feel that this is a pretty good number. After all, if it were an exam, you would have gotten an “A”! Let’s examine this a bit more closely.

If we have a production system, we need all pieces of equipment functioning. Otherwise, a delay in one piece, like the roof bolter, will prevent other equipment from moving into that workplace. As a result, the other equipment will be delayed as well. Mathematically it can be shown that the availability of a system of equipment is equal to the product of the individual availabilities of the machines. Let’s imagine that we have the following group of equipment in our production system, along with the availabilities as shown. What is the availability of this production face?

• Jumbo drill 90%
• Scaler 96%
• Rock Bolter 92%
• Wheeled loader 92%
• Mine truck #1 88%
• Mine truck #2 91%
• Feeder-breaker 98%
• Conveyor system 94%

Can you believe it? 54%! In other words, at this production face, they are mining only half the shift! And if you really want to have an anxiety attack, let me tell you that we haven’t accounted for other mining delays – unexpected rock falls or an explosives misfire, for example! We can drop down close to 40% overall availability if we aren’t careful. Clearly, there’s room for improvement – better engineering, better maintenance, and so on!

Lesson 5.2 Summary

In this lesson, we’ve learned about auxiliary operations and how they fit into the mining cycle. We’ve also gained important insights into ground control, ventilation, and maintenance operations. We’ll see how these auxiliary operations combine with the unit operations to complete a cycle of operations for the different mining methods.

5.3.1: The Continuous Mining Machine

A picture is worth a thousand words… so, let’s look at the front of a continuous miner. The cutting drum is about 2’ in diameter and perhaps 12’ wide. It is laced with carbide cutting bits, and the drum rotates at about 1 revolution per second. The cutting head is sumped into the top of the seam, and then the drum is pulled down, which is known as shearing. The sump and shear is completed in well under a minute and then repeated. The cut coal drops to the floor, where it is scooped into the belly of the machine. Look at the base of the machine – it sort of resembles a dustpan. There are two gathering arms, which move to pull the cut coal into the center of the machine.

Figure 5.3.1: Continuous Mining Machine

source: Joy/Komatsu [42]

Let’s look at the next picture to understand what happens to the coal.

Figure 5.3.2: Front end of a Continuous Mining Machine

source: Joy/Komatsu [42]

Here you can see it more clearly. The gathering arms pull the coal onto a chain conveyor, which runs the length of the machine. The chain conveyor consists of a metal trough and metal flight bars that are attached to a chain. The flight bars, which are painted black in this picture, are pulled along the metal “pan,” or trough, by the chain. In this manner, the mined material is conveyed to the rear of the machine. The rear of the machine is known as the tail or tailpiece, and it can swing over a limited arc. The tailpiece is positioned over a waiting shuttle car, and the mined coal is fed into the shuttle car. The continuous miner will continue advancing forward with the sump-shear cycles until there is a reason to stop the advance. In U.S. coal mines, the forward advance is limited by a few factors: the need to ventilate the face to maintain safe methane levels and the need to install roof bolts to prevent roof failure. Further, by law, no miners can work under unsupported roof, which will limit the amount of advance. Once the maximum advance has been achieved, the machine will be trammed back out of the cut and moved to a new face.

All right, let’s stop and catch our breath! What have we accomplished with this innovation known as the continuous miner? We’ve replaced drilling, kerf cutting, blasting, and loading with one machine, thereby reducing the number of pieces of equipment, the number of miners, and importantly, the delays. We’ve achieved a remarkable gain in both production and productivity. So much so that conventional mining of coal is no longer practiced in this country. Of course, when you make these improvements, you then expose the next “weakest link” in the system. Based on what we’ve talked about, you probably have an inkling of the weak links. Take a guess!

Many of you have probably zeroed in on the shuttle car -- when the shuttle car is filled it trams to the section dump point and another shuttle car maneuvers into position. In the time that it takes for the full car to leave, and the waiting car to maneuver into position, the continuous miner is idled. That’s a delay that is going to adversely impact our “numbers”, i.e., production and productivity. We’re going to talk more about these haulage bottlenecks shortly, but as it turns out, this is a tough one to fix. There is another weak link, which some of you have undoubtedly spotted: the need to limit the continuous miner’s advance so that the roof can be bolted and ventilation can be extended.

Allow me to introduce you to the Miner-Bolter. It took a number of years to work out the kinks, but today these machines work well. There are different configurations, but the principle is the same: incorporate the bolter into the miner. A bolter operator stands on the platform immediately behind the bolter carriage on both sides of the machine. While the cutting drum is cutting coal, the bolter operators are drilling holes and installing bolts. There is a linked platform to allow the bolters to remain stationary while the cutting head is advancing forward.

Figure 5.3.3: A Miner-Bolter

Source: Joy/Komatsu [42]

Hydraulic canopies are essential to protect the bolter operators while they are drilling and bolting. The canopies are not extended in this picture, but in the next photo they are, giving you a clearer view.

Figure 5.3.4: Miner-Bolter with Canopies Extended

source: Joy/Komatsu [42]

Typically, there will be a ventilation tube, on the order of 24” in diameter, hung near the roof against the side rib. This tubing will be extended as the continuous miner advances forward, providing the air needed to sweep the face and remove dust and gas.

Although remarkable gains have been achieved, there remains the weak link with the shuttle cars. In fact, there are a couple of options that can be considered. Before taking up this topic, however, I want to say a few more things about these continuous miners.

5.3.2: Continuous Miners for Noncoal Applications

The impetus for developing the continuous miner came from the coal mines, but their application has grown beyond coal. The motivation for using them in noncoal mines is the same: eliminate delays wherever possible. In the case of the metal/nonmetal mines, the desire is to eliminate the discrete unit operations of drilling, blasting, and loading, along with the attendant delays. Unfortunately, there is a practical limitation that cannot be overcome in many cases. It is the hardness of the ore. The carbide cutting bits must be able to penetrate and fracture the ore, and do so with some speed and without destroying the bits in a short period of time. The limit is currently a compressive strength of around 80 MPa. This means, for example, that coal, salt, trona, potash, and some copper deposits are suitable, whereas most limestone and lead/zinc deposits are not. Pushing that limit ever higher is the goal of researchers, manufacturers, and the mining companies.

The structure and appearance of the continuous miner will change somewhat for those machines designed for use in harder rock deposits. The miner shown in the picture below looks similar to the one designed for operation in a coal mine, but there are a few subtle and important differences. The size of the machine reflects the thicker deposit in which this machine will be used, the structure is reinforced, and the lacing pattern of the bits is different – larger but fewer bits to concentrate the available energy to achieve fragmentation of the harder rock.

Figure 5.3.5: A Continuous Miner for noncoal applications

Source: Joy/Komatsu [42]

As the hardness of the rock increases, the cutting head will become much smaller, but for the same reason – concentrate more energy into a small area of the rock. Look at what’s happened to the size of the cutting head on this machine. The arm, on which the cutting head is mounted, has the ability to swing and extend. Note the bit marks in the top of the deposit. This configuration of continuous miner is often known as a roadheader, which is based on its European heritage and its use in driving gateroads. Also, take note of the size of the gathering pan and the two gathering arms.

Figure 5.3.6: A Roadheader

Enter image credit here

With this overview of continuous miners behind us, we can now return to this question of the shuttle car, and what to do about the inherent delay. Well, it turns out that this is really part of a larger and age-old problem: batch versus continuous haulage.

5.3.3: Batch versus Continuous Haulage

Batch haulage is the term given to the discrete movement of material. Shuttle cars, rail cars, hoist skips, and trucks are common examples of batch haulage. Conveyors and hydraulic slurry are two examples of continuous haulage. Over the years, the goal has been and continues to be, the replacement of batch with continuous haulage. And, we’ve been reasonably successful in many cases.

Rail haulage has been replaced with belt conveyors in many mines. Moving trains around the mine, ensuring a supply of empty cars where they are needed while moving loaded cars out of the mine and doing so over limited track networks has been a daunting challenge. Gone are the days when rail haulage was used in open-pit mines and most coal mines. However, there are still applications where it is the haulage of choice. If you have to move enormous volumes over a great distance, it is a viable choice. Probably, you will find the greatest use of rail in very deep underground metal mines.

Skip haulage or hoists have been replaced in many mines by driving slopes and using rail or belt haulage, with the latter being more prevalent today. In shallower underground mines, vertical belts are being used to replace hoisting up shafts. Still, in many mines, there is no reasonable alternative to a hoist. Deep mines, thousands of feet below the surface, rely solely on skips to remove the ore. Mines at which there is no ability to put in a slope, and which are a bit too deep for vertical belts, will continue to utilize hoists.

Truck haulage continues to be prevalent in surface mining because they can transport huge volumes of material economically from the working face to the processing plant. Truck haulage is also common in underground hardrock mines that have a slope or ramp to the surface so that the material can be transported from the working face to the plant. Oversize material is generally not a problem for trucks, but is a big problem for conveyors. Abrasive rocks are not a problem for trucks, but are for conveyors.

In certain open pit mines, which are very deep, the time required for a haul truck to make its way out of the pit can become excessive. In these cases, high angle conveyors can be used to good advantage.

The haulage examples that I’ve described so far can be categorized as outby haulage, i.e., the part of the materials handling system used to transport the material through the mine and to the plant. This distinction may seem odd. In the case of truck haulage, as we just discussed, the truck is loaded at the face and then travels to the plant. However, there are many examples, particularly in underground mines, where there is an intermediate form of materials handling. In these instances, the ore at the face is loaded onto one form of materials handling, transported a short distance, and then transferred to the outby haulage system for the trip out of the mine and to the plant.

It is in the form of intermediate haulage that we encounter our old friend, the shuttle car. The LHD would be another example. And this brings us back to that question: what can we do about that weak link, the shuttle car?

The area around the active working faces is very dynamic. It is difficult to construct semi-permanent infrastructure in this area; and it is for this reason that intermediate haulage has evolved. If we are to make an improvement, we would need a conveyor belt that was mobile, i.e., it could travel around the working faces, following the continuous miners or loaders. Prototypes of such systems have been built for use in underground metal mines to replace LHDs, but they have met with little success. In the case of soft rock applications, and notably coal and salt, there are commercially available technologies to meet the need, in specific circumstances. Here is a picture of a commercially available flexible conveyor. These units can turn corners and can be piggybacked to provide a continuous path of several hundred feet between the continuous miner and the section dump point. Recall that it is at the section dump point where the coal is fed onto the main or outby conveyor belt. This technology is being used successfully in some salt mines, but has found limited application in underground coal mines. For reasons that lie beyond the scope of this discussion, they work best in shallow mines with small pillars; and that is a small subset of all underground coal mines.

Figure 5.3.7: Flexible Conveyor Train

Source: Joy/Komatsu [42]

But not to despair, there is one other option in limited but successful use.

Rather than having the continuous miner load into a shuttle car, and being captive to the availability of shuttle cars, what if we allowed the continuous miner to dump onto the mine floor? Sounds crazy, right? But, not really. We could use a loading machine to clear away the pile, and trust me, loading machines can load a shuttle car very quickly. So, the shuttle cars would queue up for loading, but if there were a delay, the continuous miner would be able to continue mining without incurring a delay. Sure, the pile behind the machine might grow rather large, but no worries. The loading machine will get rid of it quickly. As you can see in the picture, the loading machine has a larger set of gathering arms than the continuous miner. The operating principle is the same, however: the gathering arms pull the coal into the belly of the machine and onto the chain conveyor.

Figure 5.3.8: Loading Machine

Source: Joy/Komatsu [42]

And now, having addressed the quest for continuous, we not only understand the technology options, but we’ve become more familiar with the key concepts of availability, system delays, and batch versus continuous haulage.

There is one final unit operation to study, and one that is a huge part of most mining methods. It’s generally one of the more popular topics among mining students as well: explosives and blasting!

6A.1.1 Social - Economic

First, we’ll illuminate through examples what we mean by social and economic, and then we’ll look at the intersection of the two, which forms the equitable region. Similarly, we’ll use examples to illustrate that which many would consider as equitable. Please note that these examples are not exhaustive, but rather are intended to give you a deeper understanding of each term.

The social dimension of sustainability would consider the following questions.

• Will (does) the project improve the infrastructure of the community, e.g. schools, hospitals, water supply, and highways?
• Will (does) the project contribute to the local economy, through tax revenue for example, in addition to creating jobs, and is this contribution in reasonable proportion to the value obtained by the company?
• Will (does) the project provide good-paying jobs within the local community?
• Will (does) the company provide training and education opportunities to improve the skill sets of workers?
• Will it be necessary to displace or move indigenous peoples as a prerequisite to mining activity?
• What will happen to the community or region when the project is completed? Will there be other industries to employ the workers from the former project?

The economic dimension would consider the following question:

• Is (will) the financial performance¹ of the project (be) consistent with investor or company expectations?

This single question captures and represents the sum total of everything that affects the cost of bringing a mineral product to market. It also reflects market conditions, i.e. the price at which we can sell our product and the amount of product that we can sell. However, for this discussion, we will neglect market conditions and instead focus on the cost side of the equation. The mining and processing costs will be based on the many factors that we’ve studied in this course, e.g. ore grade, depth of the deposit, geotechnical characteristics of the orebody, the extent to which mechanization and automation can be applied, and so on. Of special interest here are any expenditures that would be made to address the social dimension of sustainability, such as strengthening the community through the improvement of infrastructure.

The intersection between the social and economic dimensions is aptly named equitable. Are the economic benefits that will accrue to society, and in particular the community, in reasonable proportion to the social costs of the project and to the economic benefits that the company will realize from the project? This is a difficult question to answer – how do you calculate this value? While every situation is likely to be somewhat different, there has to be genuine respect for the community and its institutions, as well as a desire by the company to improve the community within the realistic financial constraints of the project.

Unfortunately, it may become even more complicated. Whether or not a solution will be considered equitable can depend on the ethical framework under which the proposed solution is evaluated. Let’s take a non-mining example to illustrate this. Suppose that it is determined that a dam is needed at a certain location on a major river. The dam will provide flood control, sparing towns along the river from the devasting floods that occur every decade or so. The reservoir created by the dam will provide a more stable source of water for communities, and it will create some recreational opportunities as well. In total, thousands of people will benefit if this dam is built. Those are the “positives.” What about the “negatives”? There are a few dozen houses and farms that will become uninhabitable as the water accumulates behind the dam. In some cases, generations of the same families have lived in this area. The entity proposing the dam, which in this case is a government body rather than a private company, will pay the displaced landowners a substantial premium over full market value for their residences. Nonetheless, some landowners do not want to relocate and are opposing the construction of the dam. What to do...

Has an equitable solution been proposed? On the face of it, it would appear so. The landowners who will be displaced will receive sufficient money to relocate and are getting an additional sum of money for their inconvenience. Indeed, this and similar scenarios play out on a regular basis for infrastructure projects, and this is supported by the utilitarian school of ethical behavior. This school is about the greatest good for the greatest number of people. When viewed through the lens of the utilitarian ethic, the proposed project is ethical and this will strengthen the assessment that the action is equitable as well. Many industrial projects, including mineral projects, have long been evaluated under the utilitarian ethic.

In recent years, however, some have been applying another school of ethical thought known as deontology, which is concerned less with what is “good” and more with what is “right.” This is a school of thought concerned with social justice and the idea that basic human rights supersede what is good for society at large. When viewed through this lens, the dam project is unlikely to be deemed equitable, and as a consequence, there are likely to be protests, government appeals, and other actions to derail or delay the project. For mining projects, we have an obligation to address the parameters of the social and economic dimensions to achieve something that will be deemed equitable when viewed through the lens of the utilitarian ethic, and it is in our best interests to try to understand and address concerns when viewed with the deontological ethic. In essence, the evaluation of the utilitarian school is focused on rightness or wrongness of the consequences of actions, whereas the deontological school is focused on the rightness or wrongness of the actions.

¹ Financial performance is assessed through metrics such as net present value (NPV) of the project, discounted cash flow internal rate of return (DCFIRR), payback period, and earnings before interest, taxes, depreciation, and amortization (EBITDA). This will take into account many of the locational, natural and geologic, and socio-political factors discussed in Lesson 4.1.

6A.1.2 Social - Environment

Previously, we looked at a list of questions to help understand the social dimension of sustainability. Now we need to identify relevant questions focused on the environment dimension.

• Will the project adversely affect air or water quality during operations or after the project is completed?
• Will the project create other hazards during operations or after the project is completed?
• Will the project create nuisances, e.g. noxious odors or noise?
• Will the project spoil the landscape, or otherwise reduce the recreational value around the project?

Operating permits, legislation requiring reclamation as well as the Clean Air and Clean Water Acts to protect the air and water quality, guarantee that the environment dimension is well managed... except for the last question in the foregoing list.

The intersection of the social and environment dimensions is identified as bearable. This region represents a solution in which the environmental costs of a project are deemed acceptable when weighed against the social benefits of the project. As with our last discussion of the equitable region, there is no definitive quantification of “bearable,” and as such, it is subject to the interpretation of the parties affected by the project. Consequently, this will likely be interpreted through an ethical lens. The deontological ethic would assert that the environment, including the landscape, its innate beauty, and its enjoyment is a right of everyone; and therefore, regardless of any benefit, no one has the right to impinge on the landscape. Although the surface area of land that is affected by mining is extremely small, it is generally impossible to surface mine without changing the appearance of that parcel of land. It can be reclaimed, and perhaps to even better use than before, but the original appearance is likely to be changed. Indeed, this change in appearance is often an underlying motivator for protests against mining projects.

6A.1.3 Environment - Economic

We have already identified relevant questions to characterize these two dimensions. A consideration of the intersection, defining the viable region, requires consideration of more specific and technical questions, beyond those already posed. Unlike the considerations of the bearable and equitable regions, the viable region is completely definable by the engineering and science of environmental protection.

We, as engineers, define the engineering steps necessary to protect the environment, in terms of air and water quality, and also in terms of mine closure considerations. Moreover, our reclamation plan can be designed and its costs calculated. Thus, we are able to quantify the costs of protecting the environment. We can even choose to take proactive measures above and beyond those required by any regulations. Of course, that will entail an additional cost, and at some point, the cost of such measures could sink the project. Hence, using the name viable to define the intersection is quite appropriate; and if the cost to protect the environment is too great, the project will no longer be viable.

6A.1.4 Sustainable

The intersection of the three regions, bearable, equitable, and viable, defines the sweet spot of sustainability. If you think about it, how could it be anything else? The needs of society for minerals and the needs of the mining company to satisfy the expectations of their shareholders will be balanced against the constraints of the environment and the need to operate in an ethical and socially responsible fashion.

The goal of this discussion has been to equip you with an understanding of the evolving expectations for sustainable development and the ways that society views and evaluates industrial activities such as mining. No doubt you appreciate how difficult it is to establish whether or not something is bearable or equitable, and undoubtedly you can imagine how difficult it could be for a company planning a project over which some are opposed. Despite the uncertainty and fuzzy nature of bearable, equitable, and to a lesser extent viable, there are concrete actions that you can take in the planning and operations stages to facilitate sustainability. We are going to take a look at these in the next Lesson.

6A.2.1 Mining Practices

All mining, past and present, has occurred on less than a fraction of a percent of the Earth’s land mass; and the mineral resources mined to date are a small fraction of the total estimated resources. We, as miners, cannot control societal demand for minerals, but we can take steps to ensure that resources are not squandered. We do this through a design and planning process that allows us to achieve the highest extraction rate that is safely possible. In surface mining, we normally have strong economic incentives to recover all of the ore above the cutoff grade. In underground mining we usually need to leave some resource in-place for ground support reasons, and this will limit the extraction ratio. Extraction ratios of 60% are not uncommon, although for some commodities, the ratio will be much higher. Remember that after we have finished mining a specific reserve, it is almost always impossible to go back at a later time to recover additional ore. It is often said that the resource has been “sterilized.” From a sustainability perspective, we want to ensure that we are not wasting the opportunity to recover all the resource than can be recovered safely and economically; and we do this through proper planning and design.

Mines receive permits to operate, and these permits typically define conditions of operation. The constraints imposed by the permits may include operating hours, noise levels, water usage and discharge, and so on. These limitations are generally in addition to regulatory constraints, such as the clean air standards. Regardless, it is in the company’s best interest to ensure that operations are always practiced within permit limitations.

Often mines are located near or in communities, and their appearance from public areas, e.g. roads and residential developments is a concern. The use of landscaped berms to completely shield operations from public view is a good practice, as are attractive and well-maintained entrances to the mine property. Admittedly, there are some large surface mines that cannot be shielded from public view, e.g. Bingham Canyon copper mine These are the exception, and even in those cases, steps should be taken to improve the appearance through increased green areas on the site and better management of waste and product piles.

Blasting practices at surface mines located near communities are always contentious. We’ll talk about this again under community relations, but the design guidance given in Module 7 is critical to minimizing the technical and public relations problems arising from blasting.

Reclamation is of course mandated by various regulations. Often, within the conditions of the regulations and permits, there is some leeway in timing and methods. In addition to the traditional operating considerations affecting reclamation, the view and perception of the public should be factored into the reclamation planning process. Minimizing both the extent of exposed areas and the time that they remain exposed prior to reclamation becomes important if you are concerned about the bearable in addition to the viable region of the sustainability diagram. Even in operations where much of the reclamation cannot occur until near the end of the mine life, e.g. an open pit mine, small actions can be taken to think about the placement of overburden and waste piles for example. It is likely that the placement of these piles will be driven primarily by mine planning concerns, but whenever possible, sustainability considerations, balancing the viable and bearable regions, must be part of the thinking process.

6A.2.2 Community Relations

Simply put, if you want to be accepted into a community, you have to behave like a citizen and strive to become part of the community. You can improve your chances of becoming a part of the community if your behavior and actions strengthen the sense of community.

As a starting point, it may be helpful to think about a few realities. First, even if most people recognize that your mine is providing good jobs and contributing to the tax base, they still don’t want you in “their backyard.” We all drive cars, but how many of us want a car factory nearby? We all take comfort in knowing that a hospital is available in our community, but how many of us would be happy if they broke ground for a new hospital down the street from our house? You get the idea! It’s nothing personal against you or even the company, but people don’t want an increase in traffic, a noise source, or an eyesore in their community. They are concerned about their quality of life and the value of their home. Understanding their concerns is a good starting point to help you understand the task that lies ahead and the journey to become a valued member of the community.

While there is no one “magic bullet,” there are actions that many companies have found productive. Here are commonly mentioned ones.

Community Day: Invite your neighbors for a tour of the mine. Once they better understand what you do and how you do it, they are likely to have a better impression. Combine this with a cookout and a social opportunity for the community to meet and talk with the people who work at the mine. Don’t be afraid to share reclamation and mine closure plans, or to answer questions about the operation. Rather than simply complain about public ignorance, accept that it is your responsibility to educate and inform! And of course, recognize that this is something that requires attention year-round, not just during Community Day.

Citizens Advisory Group: If there are concerns and there is an active citizen's group, work with its members to form an advisory group where they can channel their concerns and engage productively with mine management. However, please understand that if you do this, you and your management must engage in good faith. Otherwise, you are likely to take a difficult situation and turn it into a very bad one!

Employee Engagement: Encourage employees to be active on the school board, in civic organizations, e.g. Rotary Club or Lion’s Club, in schools as speakers, in coaching of kids' sports, or in the scouts, among others. Adapt employee work schedules insofar as is practicable to facilitate their participation. Donate rock kits or informational literature to the schools. Invite science teachers to visit your operation.

Donate to Local Activities: Sponsor the local youth teams for soccer, baseball, softball, and so on. Buy the team new uniforms. Purchase sponsorships in local civic events. Make a donation to the local library’s annual fundraiser.

In-Kind Contributions: depending on the nature of your mining operation, you may have surface construction equipment, trucks, wheeled loaders, dozers, and so on, and you may have construction aggregates and even concrete and/or blacktop (bituminous concrete). You may be able to donate the equipment, along with operators and materials, to redo the local ballfield, or to put in or extend a local bike path, for example. If you cannot support the entire cost on your own, you might be able to donate just the machine time or perhaps the materials. These efforts can net goodwill for years to come.

Be a Good Neighbor, Every Day: Do not play “fast and loose” with the parameters of your operating permits, and make sure that contractors and customers coming onto your site do the same. If trucks are taking product from your yard, make sure they are covering their beds before leaving the property, and while you are at it, make sure they are not tracking mud or stone dust onto the public road. Undercarriage water sprays are commonly used in some operations. Back-up alarms are often noted as the single greatest aggravation within the community. When MSHA regulations allow, consider using strobes instead of audible alarms.

You can look at these community relations activities as a major investment in the equitable region of the sustainability diagram. And remember, it’s not about how much money you spend, it’s about making a genuine effort to strengthen the institutions and organizations that comprise the community through your company’s involvement.

6A.2.3 Safety, Health, and Environment

There were four mine explosions in the U.S. in the first decade of this century, including one that was the worst such explosion in the past 40 years. Globally, there several high-profile mine disasters from New Zealand to China to Brazil, and many other countries in between. Although the number of fatalities in each of these was far less than in many other disasters, the public has a much lower tolerance for a mine fatality than other disasters. A typical response to the news headlines by the average person on the street was ‘this is terrible, why do we need to mine, why should we be putting these people at risk?’ And as you might expect, these events provoked strong legislative responses across the globe, resulting in not only legislation for improvements, but in some cases, a substantial over response.

As this played out, the U.S. industry in particular, but the global industry as well, realized that they needed to be more proactive. They realized that it was not sufficient to merely comply with regulations. They needed to take additional steps to eliminate: major safety hazards, e.g. a mine explosion, that result in multiple fatalities; major health hazards, e.g. silica dust, that result in debilitating occupational diseases; and major hazards, e.g. tailing dam failure, that result in environmental disasters. During this period there was every reason to believe that one more high-profile disaster would result in the mining industry losing its “social license” to mine. In other words, the public, and by extension their legislators, would decide that ‘enough-is-enough and we’re going to ban this activity regardless of its other benefits.’ The industry was duly alarmed and knew they had to take action to eliminate these disasters².

The focus of this discussion is not the general topics of safety, health, and environment, how we apply our engineering and science skills to achieve safe and healthful workplaces, and do so in an environmentally responsible fashion, but rather on the management of safety, health, and environmental activities to help eliminate mining disasters. We, as industry in general and mining in particular, have been complying with regulations for decades, and yet every year workers die and environmental accidents occur. Safety and health were the focal points of the conversation, and the turning point was the report published by the National Commission for Mine Safety, Technology, and Training. This report advocated a more aggressive and proactive approach to mine safety, and the CEOs of the major mining companies signed a pledge committing to take the steps necessary to eliminate fatalities and reduce injuries. Mitigating environmental risks was not addressed at that time. However, the methodology to achieve this concept of zero harm³ in safety and health can be applied directly to the environment scenario as well. In the next lesson, we’ll introduce this methodology.

²In the U.S., a mining accident in which three or more persons are killed is designated as a disaster. This number is determined by MSHA and the number sometimes changes over different presidential administrations.

³ This implies that the activity should be conducted in a way that results in zero harm to the mineworker's safety or health. The concept is directly applicable to the environment as well, indicating that the activity should result in no permanent harm or damage to the environment.

6A.3.1 CoreSafety

Figure 6A.3.1. A Framework for Achieving Zero Incidents

source: National Mining Association, used with permission

Leadership is essential in affecting the behavior of workers. Safety and health performance is directly enabled (or hindered) by the behaviors and decisions of company leaders. Leaders have a responsibility for ensuring safety is integrated into all aspects of the business, holding people accountable for their responsibilities, driving a safe culture, having effective safety systems, and setting a safe example, among many others.

Many organizations that have realized substantial performance improvement have identified leadership development as the catalyst for that change. Leadership development is the process of identifying critical leadership competencies and providing structured development opportunities for leaders to improve those competencies.

While there are many leadership competencies that are complimentary to safety and health management, NMA has identified eight that are critical in the U.S. mining industry. These competencies include, but are not limited to:

• managing by objective using structured accountability;
• understanding and influencing culture;
• leveraging communication and collaboration;
• having credibility and integrity;
• using performance feedback and management;
• being action-oriented; and
• and understanding management systems and ensuring they work as designed.

Figure 6A.3.2. Illustration of the Leadership Competencies

source: National Mining Association, used with permission

Safety culture can be defined as a pattern of behavior that is encouraged or discouraged by people and by systems over time. This is a very important concept – think about what it means! Do we have a culture that values safety over production? Does the “system” reward us or penalize us if we bring safety concerns to the attention of management? Do we have established committees to look for opportunities to improve safety? We could go on with another fifty questions, but I think you are getting the idea.

Culture determines what we do when no one is watching! This statement makes it clear: if we don’t have a good safety culture, we will be unable to achieve good safety performance. The attributes of a good safety culture are illustrated in Figure 6A.3.3.

Figure 6A.3.3. Illustration of the Attributes of Safety Culture

source: National Mining Association, used with permission

The systems element of the zero-incident framework shown in Figure 6A.3.1 represents those processes and systems used to identify and address hazards. All of these are captured nicely in one overarching system known as the HSMS, i.e. the Health and Safety Management System. Often the order of health and safety is changed, so you are equally likely to see SHMS, i.e. Safety and Health Management System. The content and function of the HSMS is defined by a standard, such as the American ANSI Z10 or the Australian OHSAS 18000. CoreSafety encompasses the best of these standards. An effective HSMS will facilitate an orderly and systematic identification of hazards and the development of solutions to mitigate or eliminate risk associated with these hazards. Further, the system will facilitate the verification that the solution was effective and will incorporate regular audits to ensure that risks are managed. The flow of an effective HSMS system is illustrated in Figure 6A.3.4.

After decades of experience we know that prescriptive regulations are nominally effective, systematic control of risk offers the best opportunity to eliminate fatalities, leadership plays a critical part of improving safety performance, leadership drives culture, and culture affects systems. Figure 6A.3.1 illustrates this relationship. The integration and active management of systems, culture, and leadership can produce world-class performance. Let’s look at each of these elements in a bit more detail.

Figure 6A.3.4. The “Plan-Do-Check-Act” flow of an HSMS.

source: National Mining Association, used with permission

The circular nature of this diagram draws attention to the need for an ongoing process. Before moving on, a few additional comments are in order.

The Plan phase encompasses five activities:

• Set goals: all risks in all the company’s operations cannot be addressed at once... it is necessary to divide the work into achievable chunks.
• Define tasks: identify the task that will have to be completed. Example tasks could include conducting a site survey, reviewing past accident records, performing engineering calculations, designing risk mitigation measures, purchasing materials or hiring consultants, and installing interventions.
• Assign responsibilities: Clearly identify who is responsible for doing a task AND who will be accountable for the completion of the task. This is one of the points at which the HSMS fails because it wasn’t clear who was doing what, and then things “fall through the cracks.”
• Ensure know how: make sure the people assigned tasks have the knowledge and skills to complete the task. Although this sounds obvious, it is another common failure point within the HSMS.
• Provide resources: It usually requires time and money to implement the actions identified in the Planning Phase. Usually the money is allocated, but often the time is not. Suppose I want to implement a series of tasks for the Bedrock Quarry, and I assign responsibility for conducting the risk assessment to my quarry manager, a task that may take a few weeks to complete. I know that my quarry manager, a Penn State grad, is proficient in conducting risk analyses. Most likely, I have set up a failure at this point! Why? The quarry manager has a full-time job, from 6 in the morning until 6 in the evening, running the quarry. When will she find time to do risk analyses? She won’t. You have to allocate the resource of time if you want to be successful. Maybe we bring in an assistant quarry manager from another operation for a few weeks or find another solution.

Having completed the planning phase, we are ready to spring into action!

The Do Phase is when we conduct the work. This work could include a variety of solutions including developing an offering training, changing certain work practices, installing engineering controls, implementing different operating procedures, and so on.

The Check Phase is too easily forgotten. We identify a problem, craft a solution, and assign people to implement the solution. That’s that! We’re done! Right? Well, not so fast! Things come up, people get distracted by other tasks, and sometimes an unforeseen problem arises, which prevents full implementation of the solution. The bottom line is that someone needs to check to ensure that the solution was implemented and to provide feedback to the people responsible for the planning and doing.

Finally, there is the Act Phase. Based on our planning, doing, and checking, we have the expectation that we’ve successfully eliminated hazards and managed risks. That’s a reasonable belief to have, but it is a belief that needs to be tested. How do we “test” our belief? We assess the effectiveness of the solution. It may be something that we can visually assess through an inspection, or other instances we may be able to look at data to see whether or not a certain category of injury has decreased. Regardless, conducting audits must be a regular occurrence. Although it goes without saying that the findings of the audit must be acted on in a timely fashion, this is another potential failure point. In addition to the performance evaluations that are conducted as part of the Plan-Do-Check-Act process, progressive companies will conduct annual audits of their mines – and here’s the clever part – these audits will be conducted by the manager of a different plant or mine within the company! Nothing like a fresh pair of eyes to spot potential issues.

There is one very important piece to this brief introduction that I want to cover. Think back to the Planning phase. How do we identify risks and prioritize our work for the Plan-Do-Check-Act cycle? That is the missing piece that needs to be covered, and we’ll do that in the next lesson.

6A.4.1 Application to an Environmental Problem

Compliance with environmental regulations is an important part of the job nearly every day, and especially at surface mining sites. As with safety and health, mere compliance with the regulations is not sufficient if you are thinking in terms of sustainable development. All of the guidance given in this lesson applies to environmental as well as safety and health considerations. To illustrate this, let’s “talk through” a risk management problem related to protecting the environment, and you will soon see that the process is the same regardless of whether we are examining a safety, health, or environmental risk.

The Bedrock Quarry is located in a rural valley with no close neighbors. One neighbor of note is a large state-operated fish hatchery, which sits a few miles from the quarry. This is of note because the quarry discharges several hundred thousand gallons of water per day, and this discharge feeds into the stream feeding the hatchery. Interestingly, the fish hatchery loves the mining operation because the water they discharge is of higher quality than the normal stream water! It’s not often that a mine has a neighbor who loves them! Of course, there is a potential downside to this scenario: if the mine’s discharge should be contaminated, it could wipe out the entire hatchery. Think about the ensuing public relations disaster!

This is the kind of problem that you would tackle with your risk management tools. It is understood that you are operating with an EPA approved storm water pollution prevention plan (SWPPP), and as part of this plan, your company will have identified the day-to-day activities required to remain in compliance with the requirements of the Clean Water Act. Here we are not talking about redoing this plan or setting it aside. Instead we are going to take a fresh look to determine if we are doing everything that we reasonably can to protect the water supply. Remember, compliance alone may not be sufficient to prevent a mishap, and if we are truly mindful of staying within the bearable and viable regions of the sustainability diagram, we need to go above and beyond compliance!

As the first step of any risk management exercise, we need to put together a team. The team should include people with unique knowledge to contribute to the process, and at a minimum should include someone from operations management, an engineer, a safety or environmental staffer, and one or more miners. With the team duly constituted, there are some steps to be taken to help the individuals on the team to function as a team. These are beyond the scope of this discussion, but please realize that you can’t throw five people together, call them a team, and expect that they will function successfully as a team!

The Plan-Do-Check-Act paradigm, which was illustrated in Figure 6.3.4, provides a good road map for the team. As part of the planning, the team will define the objective, which is to ensure that the discharge water is always pure; and then they will define the tasks they believe need to be completed. The tasks will include reviewing the existing SWPPP, auditing the quarry’s performance under the existing plan, conducting a site assessment, performing a risk analysis and a risk mitigation/elimination study. The site assessment will involve several subtasks. Assignments for completing these tasks should be made and a schedule adopted. The person(s) accountable for completion of the tasks should be identified. Once these tasks have been completed, the risk management study can be conducted. The risk assessment matrix can be used to prioritize the risks that will be investigated, and then BTA or another tool of the team’s choosing, can be used to identify needed controls and recovery measures.

Once the controls and recovery measures have been identified, they need to be evaluated for practicality. Some of these may require design and construction, such as swales and rip rap lined ditches. Others may require adoption of operating procedures, such as regular inspections of fuel storage tanks and the prompt use of sorbents in the maintenance shop to cleanup spills. The monitoring and sampling program may require modifications. The “Do” stage of the paradigm, i.e. implementing the findings from the planning phase can be substantial, involving multiple personnel and requiring a period of time and significant resources to implement. This must be carefully scheduled and resourced.

All of the effort expended to reach the end of the “Do” stage will be wasted if a formal effort to verify that the risk management measures have been implemented. Moreover, the controls as implemented should be assessed to determine whether or not they are performing as expected. Of course, if the “Check” stage reveals gaps or shortcomings, they must be acted upon. This “Act” stage should not be treated as an afterthought; rather, the need to “close the loop” must be anticipated and built into the team’s charge. Adequate resources must be provided for this stage as well.

While the primary purpose of this environmental example was to illustrate the applicability of the management system approach, it should be noted for completeness that the Plan-Do-Check-Act paradigm is really an ongoing process, and the next key action would be to perform regular audits, as mentioned earlier.

8.1.1: Why do We Blast?

Yes, we blast to break the rock, but we have specific goals beyond simply achieving fragmentation of the rock! The desired outcome of a blast is broken material:

• fragmented to a specified size distribution,
• thrown a certain distance, and
• distributed in a pile that is placed to facilitate loading.

The size distribution is important for several reasons, and these reasons vary with different commodities.

Generally, the blasted material must be handled in some fashion, e.g., to load it into a haul truck or to cast it onto a spoil pile. The size of the blasted material must be consistent with the capabilities of the equipment that will be used to move it.

Oversize rock, i.e., rock that is larger than can be handled, can’t be loaded or if it can be loaded, it is too large to fit into the crusher. This creates multiple problems. It introduces delays in production, it is time-consuming, costly, and sometimes dangerous to practice what is known as secondary breakage.

The production of fines, i.e., very small particles, in the blast can be desirable in certain cases where an ore is to be beneficiated. In other cases, excess production of fines is undesirable. In the crushed stone market, for example, fines are excluded from many products, and it is expensive to re-handle and dispose of an unsaleable product.

The throw distance and the pile placement are closely related.

Pile placement

Pile placement is the distribution of the blasted material. Is the pile of blasted rock 30’ wide, 10’ long, and 10’ deep, or is 30’ wide, 80’ long, and 15” deep, or somewhere in between? There will be an optimum depth for loading. If the pile is too deep, the loader will waste time digging to load the bucket. If the pile is too dispersed, time will be wasted maneuvering over a large area to load the material.

Throw distance

This is how far the blasted material is moved. There are examples in surface mining where the blast is used to move material to a previously mined strip, rather than using a dragline or shovel. This is known as cast blasting, and obviously the distance that the blasted material is moved is a critical performance parameter. Throw is also important to ensure that the blasted material can be loaded or dug. Without adequate throw, the fragmented material will sit back down, and be extremely difficult to access. This can be a problem when advancing a face in an underground mine and, accordingly, we design the blasting pattern to ensure that the fragmented rock is lifted and thrown away from the virgin rock.

8.1.2: How do We Blast?

Very carefully! Actually, I am serious!!! There’s a difference between civil and military blasting. In the latter, their goal is usually to “blow things up.” In civil projects, such as mining, our goals are to use as little explosive as possible, as safely as possible, and to have no collateral damage, while achieving the design outcome for the blast. This is not easily accomplished and requires both good engineering and faithful implementation of the blast design by the drillers and blasters. The short answer to the question of “how do we blast” is as follows.

• We create a cavity, typically a drilled hole, to accept an explosive. The diameter of the hole will depend on the amount of explosive energy that we want in the hole.
• We will drill multiple holes, and the distances among the holes will be an engineered parameter.
• We fill the cavity, partially, with a blasting agent, i.e., an explosive. The type and formulation of explosives is an engineering decision.
• We place initiators, primers, and sometimes boosters in the hole with the blasting agent.
  • The initiators trigger the process. They can be nonelectric or electric, and some are commonly known as blasting caps. These initiators may include a time delay, which allows us to design a specific sequence for the initiation of the blast.
  • The primer is a small explosive charge that is set off by the initiator, and which has sufficient energy to set off the blasting agent. Explosives commonly used today are not cap-sensitive, meaning that the initiators do not have enough energy to set off the blasting agent. For this reason, we need to insert the initiator into the primer before dropping it into the hole.
  • Boosters are used to ensure that the explosion propagates through the entire column of explosive. This is a concern in longer holes and holes of smaller diameter.
• We connect the initiators in each hole together using nonelectric tube/cord or electric wires, and these terminate at the point where the blast will be fired, i.e., the blaster will “push the plunger” or “throw the switch.”

The placement of the holes and the timing sequence of when the holes are “fired” constitute the blasting pattern.

Collateral Damage

I said that one of the goals of blasting is to avoid collateral damage, which we do through proper design and execution of the blast. But, what do we mean by collateral damage? Here are the primary ones that we constantly have to assess.

Overbreak

This is when fractures from the blast propagate beyond the intended region. Imagine that you want to drill and blast a tunnel opening through a mountain, and you are designing for an opening that is 30’ wide and 15’ high. However, due to improper design or execution, cracks have propagated to 18’ high. Over time, it is likely that pieces of rock will begin to fall out of the tunnel roof, creating a safety hazard as well as delays in using the tunnel.

Flyrock

Flyrock is a large chunk of rock that is propelled well beyond the throw region for the blast. These chunks can weigh hundreds of pounds or even more than a ton, and travel distances of several hundred feet. Over the years, flyrock has caused numerous fatalities and millions of dollars of property and equipment damage. Proper design and execution are necessary to prevent flyrock.

Gas production

Gas production is an intended action during a blast, but the production of excess quantities of toxic gases, notably CO and NOx, is to be avoided. Improper on-site formulation of the blasting agent and problems with loading are often responsible for this problem. It should be noted that this is a hazard in both surface and underground mines.

Ground vibration

Like gas production, ground vibration is an intended consequence of blasting. However, excessive ground vibration can damage structures, and the level of vibration at the boundary of the mine’s property is regulated. Careful design of the blasting pattern, and especially the timing of the holes, is required to keep the ground vibration within prescribed limits.

Air blast

This occurs when excess energy from the blast creates a shock wave in the air. Certain weather conditions will cause the air blast to be bounced back to the surface but at some distance from the blast. There is little danger of personal or structural damage from an air blast, but it can precipitate a barrage of angry complaints from people who live in proximity to a surface mine. Nothing good ever comes from irritating the locals! Air blast can be reduced through proper design the blasting pattern and the choice of detonators used to link the holes together.

Misfires

These are not really “collateral damage” per se, but they are an unintended consequence. A misfire occurs when some of the blasting agent in a hole, or multiple holes, remain undetonated after the blast. This explosive could go off at a later time, such as during loading, and cause serious personal injury or death. Misfires can be avoided through careful attention to the execution, as well as the design, of the blast.

Good Engineering and Execution

Throughout this discussion, I have emphasized the proper design and execution of the blast. By execution, I mean the drilling and loading of the holes. Poor drilling or loading procedures will compromise the best design, and similarly, proper drilling and loading cannot compensate for a poorly engineered blast round.

It starts with good engineering, and specifically:

Design of the blasting round is concerned with:

• the choice of the blasting agent;
• the choice of detonators, primers, and boosters;
• the use of free faces. A free face is an unconfined surface, through which the blasted material is free to move unimpeded;
• the distance between holes, known as the spacing, and the distance between a hole the nearest free face, known as the burden;
• the number of holes;
• the diameter of the hole;
• the depth or length of the hole;
• the use of decking, when needed, and stemming. Decking is an inert material used at certain locations within the column of explosive when it would be undesirable to have explosive at that part of the drill hole. Stemming is an inert material placed in the top or front of the hole. It is necessary to help provide proper confinement for the blast, ensure safety, and reduce airblast;
• the timing of when each hole is detonated;
• the overall positioning or arrangement of the holes, including the use of unloaded holes;

and ends with good execution by the drillers and the blasters. Specifically for the driller:

• accurate positioning of the collar, i.e., drilling the hole exactly at the specified location;
• control the drilling angle, i.e., drilling the hole at exactly the specified angle;
• control the depth or length of the hole;
• accurate logging of the holes during drilling (applicable primarily to surface mining applications);

specifically for the blaster:

• correct mixing of the blasting agent at the hole (if batch loaded);
• proper loading procedures if there is water in the hole;
• correct loading rate;
• correct placement of decking, boosters, and primers;
• proper placement of stemming and stemming to the design depth.

8.1.3: Dance of the Detonators

I want to conclude this introduction to explosives and blasting with a short video clip (4:04) showing multiple controlled explosions. There is no narration for the video, but the blasts are accompanied with a musical soundtrack. 

Look for the following:

1. The relative timing of the holes, as indicated by when they detonate.
2. The initiation of the blast in the holes prior to the detonation of the explosive, as indicated by the flashes of light moving from hole-to-hole, as well as the small dust clouds at the collar of each hole.
3. The movement of the blasted material at the free faces (not always apparent).
4. The couple of blasts that are done underwater.

Direct link to video [60] if not showing below

I hope that you enjoyed this video. It is entertaining! It also represents good blasting practice, which you will come to better understand as you learn more about explosives and blasting practice.

8.2.1: Mixing ANFO and the Oxygen Balance

The proportion of oxidizers and fuels in the mix is critical to the performance of the explosive, and we say that the explosive must be oxygen balanced.

Oxygen Balance

It is important to achieve an oxygen balance within the explosive. This means there is exactly enough oxygen present to completely oxidize the contained fuel, but none left over to react with the contained nitrogen. We can calculate the proportion of ammonium nitrate and fuel oil to achieve an oxygen-balanced reaction. The decomposition of an oxygen-balanced ANFO is:

We can find the atomic weights for each of the elements from the periodic table.:

N = 14.006

H = 1.0078

N = 14.006

C = 12.009

And then, we can calculate the molecular weight of the ammonium nitrate and the fuel oil.

${\text{NH}}_{\text{4}}{\text{NO}}_{\text{3}}$: molecular weight = 80.05 g/mol

${\text{CH}}_2$: molecular weight = 14.03 g/mol

The ideal mixture, i.e., the oxygen-balanced mixture, consists of 3 moles of ammonium nitrate and one mole of fuel oil.

Next, we should find the weight ratio of ammonium nitrate to fuel oil.  

For 3 moles of ${\text{NH}}_{\text{4}}{\text{NO}}_{\text{3}}$ :  3*80.05 = 240.15 g

For 1 mole of ${\text{CH}}_2$:  1*14.03 = 14.03 g

The molecular weight of ANFO is, therefore: 240.15 + 14.03 = 254.18 g

The percent by weight is:

${\text{NH}}_{\text{4}}{\text{NO}}_{\text{3}}$ :  (240.15 / 254.18) *100 = 94.5% by weight

${\text{CH}}_2$ :  (14.03 / 254.18) * 100 = 5.5% by weight

So, now we know that when we mix the ammonium nitrate with the fuel oil, we need to do it in this proportion if we want an oxygen balanced reaction.

We can calculate the density of ANFO mixed to this ratio. Why we would we want to do that? If we know the weight per unit volume of the correct mix, then we can tell the blasters who make the mixture in the field to weigh one cup of the mixture to ensure the weight matches the specification weight. Note, I used “cup” as the volume measurement where one cup equals eight fluid ounces. You can use any unit, as long as you are consistent.

Let’s look at an example.

We will assume that the industrial-grade ammonium nitrate and fuel oil have the following densities: 53 lb/ft³ for the ammonium nitrate and 8.0 lb/gal for the fuel oil. How much ammonium nitrate will be required for each gallon of fuel oil to make an oxygen-balanced batch of ANFO? We will need to mix 2.6 ft³ of ammonium nitrate with that gallon of fuel oil. You should verify this result.

Continuing with this example: we know that the blasters will have a scale on the mixing truck and they will have a “cup”, which they will use to do a “cup density” check. In other words, they will fill the cup with their batch of mixed ANFO and weigh it. They will compare that weight to the weight that they have been given as the specified weight. If the blaster is using an 8-ounce container for the cup density check, what weight should be displayed on the scale?

You don’t have sufficient information to answer this question. What is the calculation that you need to do to answer the question, and what are you missing?

Although you know the weight of each ingredient in the mix, you don’t know the density of the mix. The manufacturer of the ammonium nitrate would be able to give you that information, or you could do a simple experiment in the lab to determine the density of that mix. For our purposes here, let’s assume that the density of this mix is 6.68 lb/gal. Therefore, the number that you need to give to your blaster is 0.42 lb.

The mixing equipment on the truck requires regular calibration, and even if calibrated, sometimes it can malfunction. Thus, it is recommended that blasters perform a cup-density check frequently, e.g., every 5 or so holes in a surface mine application.

We’ve looked at the ideal, i.e., oxygen-balanced mixture. Let’s look at two other cases in which we have too little or too much ammonium nitrate for the amount of fuel oil that has been added. We’ll choose a 92: 8 ratio for the first case and a 96.6: 3.4 ratio for the second case, I’ve picked these ratios so that we have an integer number of moles in our formula.

Thus, for the first case with a 92:8 ratio, we have:

And for the second case, with a 96.6 : 3.4 ratio, we have:

Let’s look at the right-hand side of the equations. In the case of ideal mix, the products of the reaction, in addition to the energy that is not shown, are water vapor, carbon dioxide and nitrogen. In the first case of too little oxidizer, i.e., the ammonium nitrate, we produce carbon monoxide, a deadly gas. In the second case of too much oxidizer, we produce NO. The gas, NO, changes to NO₂ when exposed to the atmosphere. NO₂ is a toxic, but unlike CO, which is colorless, NO₂ produces a bright yellow-orange cloud.

We are not going to calculate the volume of toxic gas produced, but be aware that it is significant. In addition to the production of toxic gas, the energy release is reduced.

The oxygen balance can be lost for varied reasons. The most common include:

• poor quality control at the batch mixing truck;
• incorrect formulation, for example by adding a sensitizer like aluminum, but not accounting for it in the addition of the fuel oil;
• water in the hole;
• lack of confinement;
• improper stemming;
• improper timing of adjacent holes, allowing the explosive to become unconfined by the time it detonates;
• open-air blasting;
• explosives at the collar of the hole (increases fume production, not fragmentation).

8.2.2: Fumes

The whole point of ensuring an oxygen balance is to achieve the maximum energy from the reaction and to prevent the generation of toxic gases. Remember, however, that the production of large quantities of gas is key to the efficacy of the explosive to fragment rock. On average 700 - 1000 l of gases/kg of explosive are produced, and these are mostly benign – N, CO₂, and water. It is impractical to achieve a perfect mix in the field, shot after shot, and consequently, a small (<4%) toxic  component of about 3% CO and 1% NO₂, depending on the oxygen balance, is not unusual. The percentage of toxic gases will increase significantly as the oxygen balance deteriorates.

Due to the possibility of a toxic component in a well-designed blast, certain precautions must be taken. The ventilation system in underground mines must be designed to dilute and carry away the gases from the blast, and no one should be allowed back into that part of the mine until sufficient time has passed to eliminate any gases from the blast. Some mines, and particularly smaller ones, make it a practice of blasting after the crew from the last shift of the day has exited the mine and the next shift won’t return until the following day. Other mines don’t have that luxury, and must ensure that a properly designed ventilation system is in place.

Fumes are normally less of a concern in a surface mine. However, it is important to note that fumes can be trapped in a pile of blasted rock, only to be liberated when the pile is loaded out. There have been a few fatalities from this, so this possibility is not to be ignored. The precaution of watering down the pile prior to loading is practiced both underground and in some surface applications. In addition to addressing a potential fume issue, the wetting action suppresses respirable dust.

A significant hazard can develop at surface mines if there is an excess production of NO₂ or CO. In the late 1990s, I was part of a team investigating a problem in the Powder River Basin, in which the large blasts were producing thick yellow clouds of NO_x – clouds that covered acres, and would drift for miles before dispersing! Fortunately, there were no reported ill effects and no fatalities, and that was lucky. My agency became involved when such clouds from a mine settled in a nearby town, near an elementary school filled with children. In the Basin, the problem was found to be water in the hole primarily, and loss of confinement as a secondary cause.

There was another illustrative problem that we investigated involving CO. These were trench-blasting applications, and there were a few fatalities due to CO poisoning. CO from the blast was trapped in the ground, but over a day or so, the gas migrated along a pipeline and entered a structure (basement), creating a toxic environment in the basement.

Fumes are an expected consequence when explosives are used. Proper procedures can help ensure an oxygen-balanced explosion, but will not guarantee a blast completely free of toxic fumes. Precautions, such as those mentioned here, must be implemented.

8.2.3: Classification of Common Industrial Explosives

The U.S. Bureau of Alcohol, Tobacco, and Firearms (ATF) regulates different aspects of explosives manufacture and use. They classify explosives according to the following definitions:

High Explosive (HE): an explosive material that can be caused to detonate with a No. 8 blasting cap when unconfined; and

Blasting Agent (BA): a mixture consisting of a fuel and oxidizer, intended for blasting but otherwise not an explosive (cannot be detonated with a No. 8 blasting cap).

HE’s that can be detonated directly with a No. 8 cap are called cap-sensitive.

BA’s that cannot be detonated directly with a No. 8 cap are called cap-insensitive or non-cap-sensitive.

Low Explosive (LE): an explosive material that can be caused to deflagrate (burn) when unconfined.

These definitions are important, as the terms and the underlying concepts are in everyday use. However, don’t worry about "what is a #8 blasting cap…" just know what it means to be cap sensitive or cap insensitive.

Examples of products in these classes are:

• High Explosives
  • Dynamites
  • Gelatins
  • Semi-gelatins
  • Water gels & slurries and Emulsions
• Blasting Agents
  • Water gels & slurries and Emulsions
  • ANFO
  • Blends
• Low Explosives
  • Black powder

Dynamites are rarely used industrially today because of safety concerns. For that matter, there is little use for the gelatins, semi-gelatins, and binaries per se in mining applications. Water gels, slurries, emulsions and ANFO blends are the predominate explosives in use. Note, however, that water gels, slurries, and emulsions can be formulated to be cap sensitive. This is why I have shown them under high explosives and blasting agents.

The use of black powder in underground coal mines was outlawed decades ago in this country because it will ignite coal dust and methane mixtures, making it an explosion hazard in these mines. Unfortunately, you will find it in use in the mines of some lesser-developed countries. The safe alternative to black powder is a permissible explosive, although there is little demand for low explosives in modern mining operations. We’ll talk a little more about this near the end of this lesson.

The blasting agents are often categorized as dry and wet blasting agents.

Dry Blasting Agents

The dry blasting agents are the ANFO blends, and are not cap sensitive. ANFO has the following characteristics:

• the ratio of industrial-grade AN to No. 2 fuel oil is 94.5 : 5.5 % by weight;
• aluminum particles can be added, up to 6 % , by weight, to increase the energy (heat) output;
• it is available in bulk (most common) or packaged;
• it has poor water resistance;
• the critical diameter is approximately 4";
• the specific gravity is over the range of 0.75 - 0.95;
• the detonation velocity is approximately 15,000 fps.

The industrial-grade AN is normally provided as prills, which are uniform beads of a few millimeters in diameter. We haven’t defined some of these characteristics, such as critical diameter, but will do so shortly.

The poor water resistance of the ANFO blends is a serious drawback because water is present more often than not. Sometimes, several feet of water will accumulate quickly in a vertical hole, and other times, a small amount of water will seep into the hole after loading. Regardless, this creates a significant problem. Wet blasting agents were developed to have better water resistance, and contain more than 5% water by weight. Another important characteristic of the wet-blasting agents is their higher density, which translates into being able to load more energy in the hole.

Wet Blasting Agents

There are two major types of wet blasting agents: water gels & slurries and emulsions. Water gels and slurries are technically different, but in common usage, the two terms are used interchangeably.

• Water gels consist of an inorganic oxidizer such as ammonium nitrate with a gelling agent along with additional suspended oxidizers, fuel, stabilizers, and so on. They are a colloidal suspension of solid AN particles suspended in a liquid AN solution that is gelled using cross-linking agents (think Jell-O), containing up to 20% water. The explosive becomes slurry with the addition of suspended solids and, in fact, slurries are more common. However, in everyday usage, the words are used interchangeably and we will do so here as well. Other key features of slurries include:
  • they can be made cap sensitive, for example, adding up to 18% Al;
  • they are available in bulk or packaged;
  • they have good water resistance;
  • the critical diameter is < 1";
  • they have poor low-temperature performance;
  • their specific gravity ranges from 1.15 - 1.45;
  • the detonation velocity ranges from 14,500 fps - 18,500 fps.
• Emulsions consist of a two-liquid phase containing microscopic droplets of aqueous nitrates of salts (chiefly AN) dispersed in fuel oil, wax, or paraffin using an emulsifying agent (think Mayonnaise). Stabilizers, and so on are added to round out the mixture. Other key features of emulsions include:
  • emulsions are more efficient than slurries;
  • they are available in bulk or packaged;
  • they have excellent water resistance;
  • the critical diameter is < 1";
  • they have poor low-temperature performance;
  • their specific gravity ranges from 1.1 - 1.3;
  • the detonation velocity ranges from 14,500 fps - 18,500 fps.

Blends are a mix of dry ANFO and emulsion, and this mix is often known as heavy ANFO. ANFO is inexpensive and emulsions are expensive. The blend is designed to capture the advantages of an emulsion, but at a lower cost. Specifically, the addition of the emulsion will improve the water resistance of straight ANFO, and it increases the density of the explosive, which means that more energy can be loaded into the hole. The ratio of ANFO is emulsion will range from 80:20 to 20:80. As the percentage of emulsion increases, the desirable characteristics of water resistance and density increase, but then, so does the cost. In practice, you would work with application engineers from the manufacturer to achieve the balance that best matched your unique needs. Other key features of blends include:

• emulsions are more efficient than slurries;
• they are available in bulk or packaged;
• they have improved water resistance;
• the critical diameter varies (<4” but greater than 1”);
• they have poor low temperature performance;
• their specific gravity ranges from 1.15 - 1.3;
• the detonation velocity ranges from 16,500 fps - 17,500 fps.

8.2.4: Permissible Explosives

The term Permissible Explosive can be traced back to early in the 20th century, when thousands of miners were being killed each year in the underground coal mines. Mine explosions caused many of those fatalities, and the mine explosions were set off by improper blasting practices. The U.S. Bureau of Mines conducted research to develop explosives that would not set off the mixtures of coal dust and methane that were typically found in the mines. Over the years, permissible came to mean two very important things: first, the explosive had been tested and certified by the USBM to meet the criteria; and second, the permissible explosive would be used according to a set of required practices. Without the latter, it is still possible to set off a mine explosion even though a certified-permissible explosive is used. Eventually, by the mid-20th century, mining laws mandated the use of permissible explosives and practices.

The migration from conventional to continuous mining practices in the coal mines has dramatically reduced the demand for permissible explosives in this country. Recently, around 900 metric tons of permissible explosives were sold in the U.S., whereas, in the mid-twentieth century, that number was closer to 60,000 metric tons! Nonetheless, it is commercially available and has application when it is necessary to blast in mines where methane could be present. Such applications would include: shooting the overlying strata for construction purposes in underground coal mines, e.g., for ventilation overcasts or increased headroom for belt drives/transfer points; and shooting large roof falls so that the rock can be loaded out and removed.

The permissible explosive must be used in accordance with the permissible practices. The key practices are as follows:

• qualified or certified persons, i.e., individuals meeting the requirements specified by the mining regulations, must conduct the blast.
• black powder, aluminum-cased detonators and safety fuses are prohibited;
• noncombustible stemming must be used;
• the minimum borehole spacing is 24” in coal and 18” in rock;
• no more than 20 boreholes are permitted to be fired in a round;
• no more than 3 lb of explosives can be loaded into the hole;
• the total delay period must be 1000 ms or less;
• the interval between delays must be at least 50 ms, but no more than 100 ms;
• the air must be tested for methane immediately before shots are fired, and the shot can be fired only if the concentration is less than 1.0 %;
• unconfined shooting is prohibited;
• all of the blasting circuits must be checked for continuity and resistance before the shot, using approved blasting galvanometer or blasting multi-meter.

8.2.5: Key Properties Affecting Choice of Explosives

There are many properties that define an explosive. Some are of most value to the engineers and scientists that formulate and test explosive products. Others are useful to mining engineers designing blast rounds. We’re going to focus on the latter, and such a list would include the following:

Properties of the explosive are key to the design of the blast round, and that will become clearer when we will talk about the design of blast rounds. However, properties of the rock are equally influential, and we’ll talk more about this in the future.

• Density: kg/m³.
• Weight strength: is a measure of the explosive energy, kcal/kg.
• Bulk strength: is the product of density and weight strength of the explosive; and is indicative of the explosive energy in the hole, kcal/m³.
• Critical diameter: the minimum diameter of the explosive column to sustain the detonation or deflagration. This will determine the size (diameter) of the hole that we drill.
• Critical density: the maximum density at which the explosion may not propagate. As holes become deeper, the concern is that the weight of the column will become so great that the density of the explosive near the bottom of the borehole will exceed the critical density.
• Sensitiveness: a measure of propagating ability. You will hear this term, but won’t use it in a calculation, although it may influence your decision to use boosters in the hole.
• Sensitivity: a measure of the energy required to initiate the explosive. For our purposes, we’re usually not interested in a quantitative measure of this property; but, rather, we need to know simply: is it or is it not cap-sensitive?
• Water resistance: a qualitative measure for our purposes.

8.3.1: Bench Blasting

Here is a sketch of a bench, showing a few boreholes, and the design variables. This bench could be in an underground or surface mine. Our goal will be to calculate values for many of the parameters shown on this diagram.

Figure 8.3.1: Sketch of a bench

First, please take a look at each of these parameters, as labeled, and make sure that you understand what’s going on. There are two terms on the diagram that we have not yet defined. They are toe and crest. The toe is the bottom edge of the bench adjacent to the vertical wall. The crest is the top edge of the vertical or nearly vertical wall. This wall is often called the highwall.

The equations for determining the burden, spacing, and stemming require that we know the blast hole diameter. So, looking at the diameter of the hole is a good first step. This diameter is defined by the diameter of the drill bit.

For benching, hole diameters typically range from a low of 3” to a high of 15”. If we’re benching in an underground mine, we’ll have hole diameters near the low end of that range, and for surface applications, we’ll be in the middle to upper end of that range. Just as a comparison, I would mention that for drifting, the holes are likely to be as small as 1-3/4” and probably no bigger than 3-4”. In many cases, you will have historical practice or the practices at similar mines to narrow your choices. Other times, you will already own equipment capable of drilling a limited range of sizes, and your design will be bounded accordingly.

Here’s an interesting chart, which is a compilation of the hole diameter and bench height at more than a hundred mine sites. As you can see, there is a rough relationship between the bench height and the diameter of drill used to create the blast hole for that bench.

Figure 8.3.2: Hole size to bench height ratio chart

Given a hole diameter, D, we can calculate the burden, B, the spacing, S, and the stemming, T. We will use Ash's to calculate these parameters. Note, as indicated earlier, and as will become more apparent with the discussion of the “K” factors, these relationships provide guidance not absolute precision.

Let’s talk about these “K” factors.

K_B = 20 for underground application and 25 for surface, assuming a standard ANFO and a rock density of approximately 2.5 g/cm³. If the rock density is significantly greater or less than 2.5, then the factor should be examined. Here’s a table of typical rock densities for commonly mined minerals. As you can see for most of them, the suggested K_B of 30 to 25 will be fine. The notable exceptions would include coal and peridotite. If you need to convert these to lb/ft³, multiply by 62.3.

If you are using an explosive, x, with a different bulk strength, the factor K_B changes as follows:

K_Bx = K_B ∗ (B_Sx)^1/2, and B_Sx is the bulk strength of explosive X.

The other two factors are constants: K_S = 1 to 1.3; and K_T =0.7.

And, that’s all there is to it!  More or less… this gives us a reasonable first approximation to laying out the pattern, or at least the burden and spacing along with the stemming. There are additional considerations, as well as sources of additional guidance.

An important metric is powder factor, because it tells us something about how efficient our blast is, and if we know typical powder factors from other similar mines, we can use that number to back calculate some of our design parameters. Powder factor is defined as the weight of the explosive used divided by either the volume or weight of fragmented rock. Its units are lb/ton or lb/yd³, or in the metric system, kg/tonne or kg/m³.

The calculation of the powder factor is straightforward. In practice, you will know how many pounds of explosive were used and you will know how many truck loads of rock were fragmented; and frankly, it is a good idea to keep close tabs on this in your operation. In the design stage, you can directly calculate the powder factor.

The explosive charge per unit length of hole, M_c, is: $\pi /4\ast \text{D}\mathrm{}$^2∗ ρ explosive ,

where D is the hole diameter and ${\rho }_{explosive}$ is the density of the explosive. The weight of the explosive in the hole is therefore the product of the charge per unit length and the charged length of the hole, L_c. Remember that the entire hole is not filled with explosive. For example, the top 1/3 or so may be stemming; and in this case, the charged length would be 2/3 of the hole length.

The volume of fragmented rock is taken to be the product of the blast area, A_b, around the hole times the length of the hole. It is defined by a rectangle with the hole in the center. The two sides of the rectangle are  $\raisebox{1ex}{$1$}\!\left/ \!\raisebox{-1ex}{$2$}\right.\text{B}+\raisebox{1ex}{$1$}\!\left/ \!\raisebox{-1ex}{$2$}\right.\text{B}$  and $\raisebox{1ex}{$1$}\!\left/ \!\raisebox{-1ex}{$2$}\right.\text{S}+\raisebox{1ex}{$1$}\!\left/ \!\raisebox{-1ex}{$2$}\right.\text{S}$ . Or simply, the area of the rectangle, A_b, is the product of the burden, B, and spacing, S. The volume of broken rock is then the area, A_b, times the length of the hole. If we want the weight of the blasted material, we need to multiply this volume by the density of the rock that will be broken by the blast.

The powder factor, PF, is:

All right, we can calculate powder factor, and we understand why it is a useful metric to compute and monitor at our operation. Earlier, I mentioned that knowing factors at similar operations could be useful to us in the design stage. How so?

If we have a known powder factor to start with, we can use that to determine a reasonable burden or spacing, which could be a good “check” on our first round of calculations. We can calculate M_C directly based on the explosive that we are going to use and the hole diameter. As a starting point, we can assume that L_c will be 70% of L. We know our bench height, so we know L. The remaining unknowns are B and S. Substituting, we have:

$\text{PF}=({\text{M}}_{\text{c}}\ast 0.7\text{L})/(\text{B}\ast \text{S}\ast \text{L})$ and simplifying,

$\text{PF}=(0.7\ast {\text{M}}_{\text{c}})/(\text{B}\ast \text{S})$ , and rearranging,

$\text{B*S}=0.7*{\text{M}}_{\text{c}}/\text{PF}$

You know M_C and PF, but not B or S. However, you can compute the product of B&S; and then, you can choose one, and solve for the other. At this point, you may be thinking: but wait a minute! How is this helping me, because I still have to choose B or S. True, but you are not doing so blindly. You could use B that you calculated earlier with Ash’s formula, plug that number into this equation, and voilà, you now have values for B and S. Or better yet, rather than using the B that you calculated, you might refer to a handbook where you can find burden-to-spacing ratios for many different scenarios. For example, such a table may list a typical S:B ratio of 1.25 for the type of mine of interest to you. Thus, for this ratio:

Substituting this into the equation,

$\text{B}\ast (1.25\ast \text{B)=0}{\text{.7}\ast \text{M}}_{\text{c}}\text{/PF}$ , and simplifying, we have: $\text{1}{\text{.25}\ast \text{B}}^{\text{2}}\text{=}0.7\ast \text{Mc/PF}$ .

Solving: $\text{B=}{(0.56\ast {\text{M}}_{\text{c}}\text{/PF)}}^{1/2}$

And so, using “real-world" data for powder factor and burden-to-spacing ratio, we are able to estimate a reasonable starting point in our design. Please remember that the constant, 0.56, in this equation is only valid for this example.

Here are a couple of other things to keep in mind. The degree of fragmentation will depend on the interrelationship among powder factor, burden, and spacing. We could, as an extreme case, choose to drill one hole of say 48” in diameter and pack in with explosives. The number might look good in terms of the powder factor, but what is likely to be the outcome?

This is why it’s useful to calculate these parameters in different ways, using not only Ash’s formulas but also using the fundamental definition of powder factor and available handbook data.

Clearly, it's a bit of an art, and you will go through iterations before “perfecting” your design. Nonetheless, there is a solid scientific approach to get you reasonably close.

8.3.2: Drift Round Blast Design

For drift rounds, we are talking about driving openings, i.e., drilling, blasting, loading, and hauling, in an underground mine. These openings are horizontal or nearly so in many cases, but can be on fairly steep grades in others. Think of advancing a tunnel through a mountain – we are driving an opening through the mountain, and depending on our needs, the tunnel may be inclined or nearly horizontal.

Although the calculations that we just covered are generally applicable, designing a blasting round for drifting is more difficult. There are two reasons for this. Primarily, it is because we only have one free face. In bench blasting, we almost always have at least two. Recall, what is happening in a bench blast. As we initiate the round, the blasted and expanding material closest to the highwall is able to move freely upward and outward in the direction of the highwall. When rock is blasted, it expands by 30% or more, so it needs somewhere to expand. Furthermore, the expanding gases from the detonation of the explosive are rapidly propelling this blasted rock. After the first row of holes has been fired and the fragmented material is moving out of the way, the next row is fired, and the round evolves in an orderly sequence based on the timing delays used in the pattern. If we don’t give the blasted material the space to expand and move, we’ll have a big problem. The fragmentation will be adversely impacted, of course, but the bigger problem is that the fragmented material will tend to stay in place, and then it will be very difficult to dig or load. This is the challenge we face in drifting, because we have only one free face.

We overcome this limitation by creating a second free face; and we do this by putting cut holes in the center of the opening that we are driving. We’ll look at these cut holes in the next lesson, but, essentially, they are unloaded holes drilled in the center of the center of the opening; and then, when this new free face is created, we set off the remainder of the holes. The blasted rock from those holes is then free to expand inward towards this new free face that we created in the center of the opening, and then outward.

We can use the same formulas for drift rounds as we did for designing a blast round for a bench. However, the patterns will be significantly different. And that’s what we really need to look at next – the patterns, i.e., hole placement and timing delays that are used in bench and drift blasting. We’ll do that in the next lesson.

Before moving on to that topic, let me finish one last detail here. I said that drift blasting was more challenging for two reasons, and we just talked about the primary reason for this. The other reason is overbreak, which is fragmentation that occurs beyond the intended space. Suppose you were blasting a tunnel with a rectangular cross-section of 30’ wide by 20’ high. Your goal will be to limit fractures in the rock to the 30’x20’ opening. It’s unlikely that you could maintain a perfect opening, but limiting the fracturing beyond those boundaries to less than several inches is doable. We aren’t as particular about this in many surface applications as we are in underground mines. Why? We touched on this earlier.

These fractures are likely to become future ground control problems, costing us time and money. The fractures or cracks are likely to destabilize over time, and if there is ground water seepage, the process will accelerate. Accordingly, we have to design a round to minimize overbreak. Unfortunately, few engineers do this for lack-of-knowledge; but it is not that difficult, and we will learn how to do it in the next lesson.

8.4.1: Patterns for Drift Rounds

The overall pattern will be based on the desired shape of the blasted opening. If we are driving headings that will be used by people and equipment to access the ore and conduct the mining operation, we will strive to create openings that will be stable over time, and be of dimensions that meet certain requirements. These requirements may be openings of a minimum height and width to allow equipment access, or the dimensions may be based on production requirements. And there could be other considerations. Rectangular openings have sharp corners that tend to concentrate stresses, whereas elliptical or circular openings do a better job distributing the stress around the opening, and as such, they are more stable. The strength of the rock and the expected life of a particular opening will influence the shape as well as the dimensions of the opening.

The cross-section shown here is typical, but regardless, the content that we are going to cover applies to other shapes as well. Notice that the cross-section is divided into named regions. This can be applied to cross-sections of almost any size and shape.

Figure 8.4.1: Cross section, divided into named regions: A. Cut, B. Stoping, C. Stoping, D. Contour, E. Lifters.

The purpose of the Cut is to create a second free face, which we introduced in the last section of this lesson. There are two general types of cuts: parallel cuts and angled cuts.

The most common type of parallel cut is known as the burn cut. Burn is the shortened name for the Viburnum Cut; so-named because it was developed for use in the lead-zinc mines found in the Viburnum trend in Missouri, back in the 1950s. Its use quickly spread to many underground mines. The burn cut consists of parallel holes, sometimes of different diameters, and always some of the holes are unloaded. The unloaded holes within the burn are referred to as relief holes. Sometimes, a few angled holes are placed within the burn, but this shouldn’t be confused with the other major type of cut, i.e., the angled cut.

Here are examples of burn cuts. The loaded holes are shaded black. The numbers adjacent to the hole refer to the firing sequence. The spacing between the holes is shown for most of them, although the diameter is not. The diameter of the holes would be the same as used for the other regions of the pattern, with the possible exception that the unloaded holes are occasionally of larger diameter. For example, if the loaded holes are around 2”, the unloaded holes of the burn cut may be 4”.

The unloaded holes provide an initial, although quite small, free face. The holes labeled “1” are fired first. The expanding rock moves toward the center and also out in front of the free face. The result is an expanded second free face, and the holes labeled “2” are then fired. They have the benefit of the larger free face, and after they have done their job, there is a sufficiently large free face in the center for the other holes to be fired.

Figure 8.4.2: Typical burn cuts used in underground mining

Here is an example of the most widely used angled cut, which is known as the V-Cut. All of the holes are loaded with explosive, which is a contrast to the burn cut. The V-Cut shown here includes a baby V, which is also fairly common. The number adjacent to the holes designate the millisecond (ms) delay for the firing of that hole.

Figure 8.4.3: The V cut. This diagram illustrates a typical drift round in medium-hard rock using four-hole vertical V cut with two-hole baby V cut and MS delay electric blasting caps.

In this example of the V-cut, the baby V fires first, and creates a small free face for the firing of the two large Vs. Then, after the two large Vs have fired, there will be a sufficiently large free face to fire the remainder of the holes. Other variations of angled cuts include pyramidal and diamond-shaped arrangements.

A particular cut will tend to give better and more consistent results in a given deposit. Most underground limestone mines, for example, will settle on a V-cut after trying the burn and V-cuts, whereas in the lead mines in Missouri, there is a decided preference for the burn cut. One of the more important deciding factors will be the pull that can be achieved with the round. Pull refers to how far back into the face you can successfully blast. If we drill a hole, say 12’ back into the face, can we get good fragmentation at the back end of the hole? As you attempt to drill further and further back into the face, it becomes increasingly difficult to achieve good fragmentation near the back of the hole. The distance that you can achieve good fragmentation is known as the pull. It’s important to monitor the pull closely because if a problem has crept into your workflow, whether it is a change in the geology, an inexperienced driller or blaster, or other problem, it will often show up first as a reduced pull.

The stoping holes do most of the work of advancing the face, as they blast the most rock. These holes will begin to fire as soon as the cut holes have cleared, and they will fire in a sequence such that the holes closest to the cut will fire first, and then those further away will fire in sequence. The idea behind the sequencing of the stoping holes is to ensure a good free face for the blasted material to move through.

Lifters are not always used, but their purpose is to help ensure good fragmentation at the bottom of the drift, as this will allow for an even surface that will facilitate movement of people and equipment through the drift. Lifters can also help to throw the blasted material out into the drift to improve the ease of loading.

The holes in the Contour region are designed to ensure a smooth surface and to minimize overbreak. The contour region is properly defined by two rows of holes. The inner row is known as the buffer row, and the outer row at the perimeter of the opening is known as the contour row. Research completed in recent years provides design guidance on how to minimize overbreak using the buffer and contour rows. However, this has yet to be widely adopted.

The delay sequence for a drift heading using a burn cut is shown below. Note that the burn has two angled holes; and also note that the contour region only has one row of holes around the perimeter, reflecting the longstanding practice.

Figure 8.4.4: A typical 10-foot by 10-foot heading with the delay sequence for each hole.

The process of minimizing overbreak is referred to as controlled blasting and is documented in the following publication: NIOSH Report of Investigations 9691, A New Perimeter Control Blast Design Concept For Underground Metal/Nonmetal Drifting Applications [63], Stephen R. Iverson, William A. Hustrulid, and Jeffrey C. Johnson.

We are not going to go through the details of the design in class, but I want you to understand conceptually the process. You have the publication that you can use in the future to design a controlled blast.

This approach requires that significantly more holes be drilled and loaded than in the traditional approach. However, it is rarely about first costs only! You have to look at costs over time, and then make an informed decision. While the first costs of this approach are higher, the improved ground control should yield larger savings over time by eliminating the extensive damage zone into the back and walls.

This approach allows a burden to be calculated between the buffer row and contour row, such that the damage radius from the buffer row does not extend beyond the contour row.

Figure 8.4.5:

The spacing of the buffer and contour holes is based on the damage radius, Rd, as shown here.

Figure 8.4.6:

After the buffer and contour holes have been located, the lifters are added. Next, the cut holes are placed, and then, finally, the stoping holes are located. Design guidance for these holes is given in the report.

It is important to note that the contour holes are very lightly loaded. The function of these holes is to trim the wall at that point, not to create additional crushing and fracturing. This is known as decoupling of those holes. There are various ways to load these holes to achieve the desired effect. One way is to use a detonating cord in the hole instead of a blasting agent such as an emulsion or ANFO. Another is to use trim cartridges, i.e., products specifically developed for this purpose.

I think I’ll close this discussion on blast rounds for drifting by showing you a short video clip. The blast will be repeated several times. You’ll want to try to determine the type of pattern that is being used, based on your observation of the detonating cord.

8.4.2: Patterns for Bench Blasting

There is a seemingly endless variety of patterns, and to make sense of it all, it’s helpful to remember that when all is said and done, the design goal is to achieve uniform fragmentation within a certain size range, and to do this it is necessary to ensure an adequate free face. Most of the variation that you will see in the patterns, and especially for surface mining, will be as result of the way in which the free faces are created and managed. The patterns for an underground bench blast tend to be straightforward – the amount of explosive that will be set off will be far more limited than in almost all surface mine blasts. As you attempt to blast larger and larger volumes, you need more holes, more rows, and the timing delays become more involved.

Here is a common surface mine layout with one free face. The numbers underneath the hole correspond to the delay sequence, first, second, third, and so forth. The numbers over the holes represent the millisecond delay for that hole.

Figure 8.4.7: Common surface mine layout with one free face. The No. 1 through No. 8 periods (25 MS through 200 MS) provide 25 milliseconds between each period; the No. 8 through No. 15 (200 mS through 500 MS) provide 50 milliseconds between each period; and No. 15 through No. 19 (500 MS through 1,000 MS), 100 milliseconds. Even with this much additional time between the rows, the tendency still exists for the rock to stack if the number of rows is excessive.

If there are two free faces, as in this case, a pattern like this is common.

Figure 8.4.8: Common surface mine layout with two free faces. When the blast area is open on two adjacent sides in an external corner (lower right above), the blast should be designed to utilize this additional relief.

Occasionally, it is necessary to design a surface blast where there is no free face. Imagine an open field, and you are going to make the first blast so that material can be removed to initiate the surface mining cycle. By the way, the name given to this is box cut. Box cuts are a common means of accessing an orebody for underground mining, when the deposit is within a few hundred feet of the surface; and they are required as one of the initial development steps in most surface mining operations. A blasting pattern for a box cut is shown here.

Figure 8.4.9: Blasting pattern for a box cut. Sinking blasts like this vary from most blasts because there is no open face for relief and movement must be vertical. Since entire blast is "in the tight," the vibration and fly rock damage potential is high.

These patterns can and will be modified ad nauseam to account for local geological anomalies, nearby structures such as pipelines or buildings and so on. With the few figures that we’ve examined here, you really do have comprehensive starting point for the design of your own pattern. You know how to: determine the burden, spacing, and stemming; choose an explosive or blasting agent; and lay out a pattern. One glaring gap in your knowledge base is how to achieve the desired timing delays. And so, we need to talk more about initiating the blast and the technology options available to us. We’ll do that in the next lesson.

8.5.1: Initiation of the Blast

Initiators contain high explosives that will detonate upon receiving an appropriate “signal.” If the blasting agent is cap-sensitive, then this action will set off the explosion. If it is not cap-sensitive, then the initiator will set off a primer, which will set off the blasting agent. In a typical pattern, there are many holes and each will have an initiator, and all of these holes need to be connected together, and a line must then be run to a safe location where the blaster will “throw the switch.”

Let’s talk about these initiators, their connection back to the blaster, and the method that the blaster uses to start the chain reaction.

Initiators can be classified into three categories. The categories and their primary characteristics are as follows:

• non-electric (75% of market and decreasing):
  • immune to stray electricity,
  • water is less of a problem than with other types,
  • modest cost advantage, i.e., lower cost than other types;
• electronic (20% of market and increasing):
  • allows for precise timing and sequencing,
  • enhanced site security,
  • immune to stray electricity,
  • higher first cost and more complicated to use;
• electric (5% of market and decreasing):
  • historically offered better control of timing and sequencing over non-electrics,
  • simplicity in use,
  • lack of site security,
  • susceptible to stray electricity.

Non-Electric Initiation

The predominate non-electric system uses shock tube and a non-electric blasting cap. Shock tube is plastic tube with a thin coating of a high explosive (think of a plastic drinking straw in which the inner wall of the straw has a thin coating of high explosive). Specially made plastic block connectors are used to join different lines (shock tubes) together; and, if desired, delays can be inserted at these connector blocks as well. Shock tube is set off with a machine that creates a spark or a mechanical device such as a spring-loaded hammer that strikes a starter resembling the tiny button primer at the base of a shotgun shell. Today, shock tube is used almost exclusively within the non-electric category. I don’t want to unnecessarily confuse the issue, but be aware that the nonelectric blasting caps used with shock tube can contain a pyrotechnic or an electronic time-delay mechanism within them. We’ll discuss blasting caps, both non-electric and electric, shortly.

Detonating cord contains a high explosive, such as PETN. It was used for trunk lines and downlines. Lines can be simply knotted together. It is initiated with a blasting cap (electric or non­electric). The detonation velocity of PETN is 27,000 fps, and it creates a huge air blast. This is a big problem, and as a result, detonating cord has been largely replaced with shock tube.

There are two other types that no longer used in modern operations, but it's interesting to know what they are in case you read or hear about them. They are safety fuse and igniter cord. Allow me to tell you a bit more about safety fuse.

• Safety Fuse is a thin cloth fiber cord with a black-powder core. It has a uniform burn rate, for instance120 sec/yd. By varying the length, you could vary the time delay. Safety fuse was used to set off dynamite. It was also used in a lot of bad old cowboy movies in which a bad guy tied the good guy to a powder keg, and the powder keg was connected to a very long safety fuse. The bad guy lit the safety fuse and rode away. Since the safety fuse burns so slowly, the viewers got to watch it burn for a few minutes, giving the hero just enough time to swoop in and extinguish the safety fuse before it set off the powder keg!

Electronic Initiation

The popularity of electronic initiation is increasing because this system offers significant performance and safety advantages over other systems.

• Electronic caps consist of an electronic chip, a small capacitor, and a high explosive to detonate a primer or a cap sensitive explosive.
• Each electronic cap has a unique serial number with a bar code.
• Electronic caps are wired into the blasting circuit with insulated copper wire.
• An electronic blasting machine is required, and it communicates digitally with each cap.
• The electronic blasting machine is programmed to send a “fire” signal to each cap, based on the cap's serial number, according to a user-determined firing pattern.
• The wired circuit connecting the electronic caps together and to the blasting machine also carries an analog voltage to charge the capacitor in each cap. Charging of the capacitor begins after the blasting machine has “authorized” the cap to fire. The capacitor provides the energy to detonate the high explosive within the cap assembly. The authorization to each cap is possible because the unique serial number of each cap is known. The blaster can use a tag reader to record the serial number of each cap as each hole is loaded. This data can then be downloaded to the blasting machine.
• The time delay is determined electronically on a chip, and this is known as an electronic delay.
• The more precise timing and sequencing can be used to minimize ground vibration as well as to improve the outcome of the blast…but only if the underlying shot design is correct. The improved blasting performance available with an electronic initiation system will offset the higher initial cost of the caps.
• This system affords the user with a high degree of site security – a cap cannot be set off without a special machine and knowledge of the cap’s serial number.

It should be noted here that there is a type of electronic blasting cap that is made for use with shock tube, and that cap offers the advantage of more precise timing, but it does not offer the other benefits inherent to an electronic ignition system.

Electric Initiation

For decades, electric blasting caps were the only choice if you wanted to control more exactly the time delays. However, their susceptibility to stray electricity (think cell phones, two-way radios, and approaching electrical storms) has led to their rapid decline.

• Electric blasting caps have a bridge wire inside of them, embedded in a high explosive. When current from the blasting machine, or stray electricity, flows in the network of copper wires, the bridge wire melts, igniting a column of powder that controls the delay time, which then sets off a high-explosive base charge. This is known as pyrotechnic ignition. The base charge sets off a cap-sensitive blasting agent or a primer.
• The blasting circuit consists of many wires connected in series and in parallel, e.g., leg wires and trunk wires. A circuit analysis must be performed to ensure that sufficient current will be available to ensure a reliable blast.

The following diagram can be very helpful to solidify your understanding. The two different delay options, i.e., pyrotechnic and electronic, can show up in different initiation systems. That is sometimes confusing when you are learning this for the first time. Thus, it really is helpful to examine this diagram until you understand the operating principles and why they are named the way they are.

Figure 8.5.1: Two delay options: pyrotechnic and electronic

Historically, time delays were used to control the orderly movement of material through the free faces, as we have discussed; or to limit the amount of explosive being fired at any one time, which reduces the level of ground vibration. The advent of electronic initiation ushered in a new era in which we could more effectively take advantage of the physics.

Traditionally, time delays were available in 25 ms delays, which is adequate for the purposes that time delays have been used. The p and s vibration waves are shown in this figure for two cases: the first using traditional delays; and the second, using electronic delays that allow timing down to 2 ms. This allows us to create a positive interaction with the shock waves, and this gives us improved fragmentation.

Figure 8.5.2: Hole delay timing

It can be shown that the time delay, T, required between holes in a row is:

where:

T = the time delay between holes in a row, ms

S = distance between holes in a row, ft

Vp = compression or sonic wave velocity, ft/s

The row spacing distance used in the equation is the effective spacing, as shown for three different cases here. Note that the definition of a “row” is based on the firing sequence.

Figure 8.5.3: Spacing by effective row distance

The value for Vp can be estimated from tables or by seeking advice from geophysical consultants. The value for Vp not only depends on the type of rock, but also the rock mass characterization. There are a few different characterization schemes in use, such as the Rock Quality Designation (RQD) or the Rock Mass Rating (RMR), but, in general, they take into account the number and spacing of the joints, the competency of the rock, and other characteristics of the rock mass. If you want a number for Vp so that you play with the equation, use 10,000 fps.

The potential for using electronic initiators to take advantage of the wave interaction is just being realized, and will become more common in the future. You now have knowledge of this, and can choose to take advantage of it in your future work.

Earlier in this lesson, I mentioned the idea of limiting the amount of explosive detonated in a given delay interval to lessen the impact of ground vibration. I’d like to look at that in more detail at this time.

8.5.2: Limiting Ground Vibration - Scaled Distance

Humans are very sensitive to vibrations and can perceive extremely low levels; and while this may have been an invaluable survival skill a long time ago, today it creates problems for mining and construction companies wanting to conduct blasting as part of their operations. Please don’t misunderstand me: I am not suggesting that we should blast in a way that causes any property damage. Unfortunately, the difference between the level at which people “feel” vibration and the level at which damage occurs is large. Consequently, when people feel vibration, they assume the worst. They look around the house and see a picture on the wall that is no longer straight. Had to be those vibrations! They look more closely, and they see some fine cracks in the plaster on the wall. Well, that’s it; we’ll have no more of this, they’re going to bring down our house!!! Then the angry phone calls, letters, and confrontations with company officials begin. Sometimes, citizens groups emerge, and usually local and sometimes state officials become involved. Mining companies spend more time and money dealing with this than they should! Oh, and when I say mining companies, I mean YOU. If you go to work in a surface mine, chances are good that in your first few years on the job, you will deal with this. So, what’s a person to do?

I have five recommendations.

• First, and foremost, you have to get the engineering right! You have to design your blasts correctly. We’re going to learn how to ensure that the ground vibration levels from your blast are below levels that could create any damage.
• Second, you have to ensure that the drillers and blasters are executing the blast as you designed it.
• Third, you need to be “smart” about when you blast. Most surface mines that are in close proximity to neighborhoods blast once a week. You’re likely to create fewer problems with the neighbors if you blast late morning or early afternoon on a weekday. Blasting around dinner time is a bad idea, as is shooting a round on the weekend!
• Fourth, you and your company should work to have a genuine and positive relationship with the community, as we discussed way back in Module 1. That will help you to navigate these kinds of issues, and, moreover, you may need to spend some time educating the locals on this topic. In fairness to them, they feel the vibrations and they are worried that you are going to damage their house. If they were structural or mining engineers, they wouldn’t be concerned, but they are not, and they are concerned! Educate and inform!
• Fifth, you should give serious consideration to contracting with a consulting geophysical firm that specializes in blasting vibration. They will help you to set up a seismic monitoring program and provide other guidance. Sometimes, neighbors will file damage claims, or raise regulatory issues; and this will happen even though you are certain that the vibration levels were well below any damage threshold. Obviously, you want to ensure that you are not actually causing a problem, but beyond that, you’ll want to establish a paper trail to document that you’ve been doing it properly, and that you could not have caused the loss that is being claimed. And, that's where a geophysical consultant can be a big help to you. There is a company based here in Pennsylvania known as Geosonics / VibraTech. They provide services internationally and are active with our Department and our chapter of SME. I recommend that you peruse their website to get a better feel for the type of services that mining companies use. Here is the link: GeoSonics/Vibratech [64]

Now, I want to focus on the engineering.

Ground vibration at a remote location is a function of the amount of explosive set off within a short time period (a few ms) and the characteristics of the rock mass between the blast holes and the remote location. You cannot change the local geology, and if you are mining stone, for example, it is a safe bet that the rock in your quarry is solidly coupled with the surrounding community; and as a result, any vibration that you create will travel quite well through that bed of limestone to the surrounding neighborhoods.

You can’t change the rock mass. That leaves the amount of explosive fired in a short time as the parameter that you can control. An explosive works by creating shock waves and gas pressure. The more explosive that you set off, the larger the shock wave, which travels through the coupled rock mass, and causes ground vibration. So, by limiting the amount of explosive set off at any one time, you can limit the shock wave. Within a few thousandths of a second, the shock wave will have dissipated, and then you can set off the next group of holes. Earlier, we talked about using time delays to create and effectively use the free face(s). Now, we’re adding a second reason for using time delays in a blast round -- to limit ground vibration (and also airblast).

We characterize the shock wave by its peak particle velocity (ppv) and the frequency of the wave. Both are important, but most of the attention is given to the peak particle velocity. Rattling of windows and dishes, or cracks in a wall can be related to the peak particle velocity. The design question for us is, how much explosive can we fire at one time without exceeding the threshold ppv value at the nearest structure?

We could model the rock mass and use sophisticated geophysical models to predict the level, and there are a few instances where such an effort would be warranted. For example, if the nearest structure is a building that has a piece of multimillion-dollar lab equipment that is extremely sensitive to low levels of vibration, then, we’re going to call in expert consultants, e.g., Geosonics, to perform the analyses. Fortunately, in most cases, we can use engineering approximations. These are normally very conservative, and after we gain some experience with the pattern, we may be able to increase the amount of explosive per time delay with no adverse impact on neighboring structures.

The U.S. Bureau of Mines developed an approximation method, known as scaled distance, to estimate the amount (weight) of explosive that would produce a certain ppv at a specified distance. Over time, this method has been adopted by state and federal regulatory agencies, and these agencies have regulations that require that ground vibration be limited based on either the scaled distance or the peak particle velocity of the shock wave. Usually, you would use the scaled distance to calculate the amount (weight) of explosive that can be fired without exceeding the statutory limit, which is typically a peak particle velocity of 0.5 in/s.

The Bureau of Mines (USBM) conducted hundreds of tests at different surface mines. The sites and surrounding areas were heavily instrumented to record the amplitude and frequency of the vibrations from blasts of varying size. Eventually, they developed a relationship between the peak particle velocity and a quantity they defined as scaled distance.

The scaled distance, S, is a function of the distance to the nearest structure, D, and the charge weight, w, fired within an 8 ms interval.

The units for D and w are ft and lb respectively. The unit for scaled distance, S, is therefore $\text{ft/}\sqrt{\text{w}}$

Rearranging the equation to solve for the charge weight: ${\text{w=(D/S)}}^{\text{2}}$

If you know the distance to the nearest structure and the scaled distance, you can calculate the charge weight that can be detonated in one delay period without exceeding a particular vibration level. So, let’s look at a plot describing the relationship.

This is a plot of the USBM data points, and each point is defined by the measured peak particle velocity (ppv) and the scaled distance. A good correlation exists as shown. They are actually plotting data from three different geophones, i.e., seismic sensors. The line of interest to us is the one labeled “vertical” because it represents the ground vibration level of interest to us at the structure we are trying to protect.

Figure 8.5.4: Plot to describe the relationship...

source: Siskind et al. 1980

To continue with our example, let’s say that the nearest structure, which we do not own, is 2500’ from the bench that we are going to blast. The statutory limit for this mine is 0.5 in/s. What is the maximum charge weight per delay period?

Maximum charge weight/delay period: ${\text{w=(D/S)}}^{\text{2}}$

${\text{w=(2500/42)}}^{\text{2}}\text{=3543lb}\text{. }$

The computed value for w is the total weight of explosive that can be detonated within a prescribed delay period, which is typically 25ms. Based on the amount of powder in each hole, you can then calculate how many holes can be fired at one time. Suppose, for example, that each hole is 5” in diameter, 45’ deep, and requires 10’ of stemming; and you are using ANFO with a density of ${\text{75lb/ft}}^{\text{3}}$ . How many holes could you fire at one time, without exceeding the prescribed vibration limit?

The charge weight per foot of hole = ${\text{(π/4)}\ast \text{(5/12)2}\ast \text{1ft}\ast \text{75lb/ft}}^{\text{3}}\text{=10}\text{.2lb/ft}$

The charge weight per hole = $\text{10}\text{.2lb/ft}\ast \text{(45-10)=358lb/hole}$

The maximum number of holes per delay period = $\text{(3543lb)/(358lb/hole)}$

= 9.9 holes.

How accurate is this estimate? We need to understand the assumptions underlying the scaled distance chart. The chart is based on the use of a dry ANFO. If a more energetic blasting agent or explosive is used, the chart will overestimate the amount of powder that can be fired at one time. The USBM field sites were predominantly surface coal mines and some quarries. Depending on the local geology, the chart will overestimate the amount of powder. An overestimation of the amount of powder translates directly to an increase in the ground vibration level, which could put you in violation of state-regulated maximum levels. So, what good is this approximation if it can lead to significant overestimates? Simply, it provides a good starting point for your design. Further, in Pennsylvania for example, the state agency requires you to start with a scaled distance of 90. It is unlikely that you would ever have to worry about an overestimation at that distance!

Our discussion has focused exclusively on ground vibration. A similar approach is used to limit airblast. A different chart is used becuase the relationship between air blast and distance is cubic rather than squared. If we were to use the companion chart for airblast, we could go through an almost identical process to determine how many holes could be fired within a delay period without exceeding the airblast threshold. We’re not going spend any time doing that, but if you have the need at a future time, I am sure you could do that without any problem!

OK, let’s recap the engineering design of the blast round, as described in the recent lessons.

1. The design is guided but not fully defined by science. Practical experience is invaluable in optimizing the design. If you don’t have that experience, the people who sell blasting products often have a wealth of practical knowledge and are happy to share it with you. Of course, they will be hoping for your business!
2. The design process is iterative. You start out assuming certain parameters and calculating others. You compare the numbers that you are generating with handbook or literature values. Then, you rework some of those initial calculations with modified values, which generates results more aligned with the practical experiences of others. In particular, you will likely start by establishing values for the hole diameter, the burden, spacing, and stemming. Likely, you will locate typical powder factors and burden to spacing ratios in similar applications. You may then “fine-tune” your burden and spacing values, and perhaps reevaluate the diameter of your borehole.
3. You will lay out the pattern of the holes based on free-face requirements, and you will assign holes to specific delay intervals.
4. If you are blasting in proximity of structures that must be protected, you will assess ground vibration and airblast potential. It is likely that you will then alter your timing to limit the number of holes that fire within a given delay period.
5. A reasonable next step would be to drill some holes and evaluate your design, keeping in mind that even if you are completely satisfied with the outcome (highly unlikely), you should be evaluating the outcome of your blasts on a regular basis, day-after-day, week-after-week, and month-after-month!
6. We saw earlier in this lesson that you can use the precise timing of electronic delays to further improve fragmentation, and probably with less explosive. If you were going to pursue that in your design, you would then calculate the appropriate delay between holes in the row. Keep in mind that these delays would all occur within the large delay period that is based on managing your free faces and limiting ground vibration or airblast.
7. There will be other reasons why you will need to modify your design over time. If you are experiencing fragmentation issues with cap rock, for example, you will utilize a set of design criteria to resolve them. If there is a big mud channel or fault running through your property, it will necessitate serious changes in your design. Those details are well beyond the scope of this course, but with your engineering fundamentals background and the knowledge that you’ve gained here, you’ll be able to handle those problems as they arise.

In the next module, I am going to show you photos taken at mine sites around the country to illustrate many of the basic concepts that you’ve studied in this module.

9.1.1: Mine Planning Details

Mine operators want to maximize profit while ensuring safe and environmentally responsible operations. There are many constraints that will limit the profitability of a mining project. Many of these constraints are not well known at a given point in time, and some of them can change unpredictably over time. These uncertainties create risk, and this creates complexities for mine operators. Major uncertainties for a mining project are grade, tonnage, and geotechnical conditions, as well as economic-related uncertainties such as fluctuations in price and demand for the mined product.

Good mine planning throughout the life of the mine is an essential prerequisite to realizing the financial goals of the project. Moreover, it is very important for you to appreciate the time value of money. Simply put, the value of $1 today is much higher than the value of $1 that will be received in the future. We use a metric known as the net present value, which allows us to compare future revenues or costs at the same point in time. Suppose, for example, that your project will produce $100 of revenue this year, $500 in year 5, and $1000 in year 10. My project will produce $600 this year, $600 in year 5 and $100 in year 10. Let’s say that each project requires you to make the same investment today, to realize these revenue streams. For this illustration, the amount of your investment is immaterial, but it is the same for both of them. Which project is the better investment? On the face of it, your project appears to be the better investment. Over ten years, it will return $1600, whereas my project will return only $1200.

However, this comparison neglects the time value of money. If I have money in hand, I can put it in the bank and earn interest, for example. Let’s suppose that the interest rate is 10%. Then, the question becomes what amount of money would I have to invest today at 10% so that in 5 years it would equal $500. In other words, what is the present value that will accumulate to $500 if it earns 10% interest each year for 5 years? The answer is $310. In other words, that future earning of $500 is only worth $310 today. What about the $1000 that you expect to earn in 10 years? It has a present value of $385. So, the present value of your project is ($90+$310+$385), or $850. The present value of my project is ($545+$372+$38), or $955. Thus, the value of my project is better than yours, based on the net present value. The interest rate that we use for this calculation is also known as the discount rate. As you can see, the present value of a dollar earned far in the future is very little. Further, who knows what the world will look like in 10, 20, or 30 years? So much can change. For these reasons, the decision to move forward with a project is influenced heavily by what happens in the first five or so years of operation. For now, do not worry about the calculation of present value, but only the concept of present value and how it influences mine planning and investment decisions.

In MNG 412, you will learn about several economic tools to evaluate the merits of a project, and net present value is only one. Payback period, discounted cash flow, internal rate of return, among a few others, will help to inform a decision on the worthiness of a given project. Regardless, without accurate mine planning, the investment calculations will be of little value!

As a general rule, if we are to maximize the project’s profitability, we need to maximize its net present value. Based on this concept, it makes sense that we plan our mining activity to recover the part of deposit that will yield the highest profits earliest in the life of the mine. Profit is determined by subtracting cost from income:

Profit = Income – Cost

In today's mining industry, it is the income or cash flow generated in the first 5 to 7 years of the mine life that will either make or break a surface mining operation, not the remote economics of the ultimate pit limit, which is usually at least 20 years away. Therefore, if we want to maximize profit, we have to define two objectives: (1) maximize the income and (2) minimize the cost.

How do we maximize income?

• by mining out the highest grade of material as early as we can
• by mining out as much material as we can sell at the earliest possible time

What are some of the ways to minimize the cost?

• minimizing waste removal operation
• minimizing the cost of the unit and auxiliary operations

A question that may rise in your mind is: Why don’t we use a couple of huge pieces of equipment to extract all of the ore during the first year, instead of spending say 15-20 years in one operation? Actually, that is a good question at this point. Here are some reasons why that strategy isn’t practical.

Insufficient Market Demand

It is unlikely that we would be able to sell all of the product in a short time. Therefore, we would incur expenses in the present year, but not receive income to offset those expenses until a future time. Given the time-value-of-money discussion that we’ve had, we certainly don’t want this. Moreover, when you have product to sell, but no buyer, you incur additional “penalties.”  Namely, you are taking up space around your plant, or you are incurring a charge to “store” the material elsewhere; and the quality of your product may deteriorate as it lies around.

Increased Equipment Costs

The purchase price of equipment increases with size and can become disproportionately more expensive if the piece of equipment is not in a common market range. Therefore, a small fleet of very large equipment will cost us much more than a larger fleet of smaller equipment of common size with the same overall capacity. However, the unit operating cost, which is the sum of maintenance and operating expenses for a fleet of very large equipment can often be significantly lower than a large fleet of small equipment.

Years ago, mining companies believed that bigger equipment is always better. Their reasoning was that even though the ownership cost is greater for bigger equipment, the unit operating cost is lower. Therefore, the overall unit cost, which is the sum of the unit ownership and unit operating costs, of the bigger equipment will be lower. It was true, to some extent. However, if a single piece of large equipment breaks down for a couple of hours, it will significantly delay production. On the other hand, if we have a fleet of smaller pieces of equipment, even if one needs maintenance, the rest of the fleet can generally cover its absence. Moreover, the unit operating cost of bigger equipment, especially the maintenance cost, dramatically increases as the equipment ages. Consequently, the “bigger is better” strategy must be carefully analyzed.

Higher Mine Development Costs

If we want to use really big equipment, then we’ll have to change the design parameters of our mine. Larger equipment requires sufficient space to operate. For example, if we’re talking about a larger truck, the haul roads will need to be wider, and we’ll need a better road bed and bearing surface, i.e., a more expensive haul road that will remain stable under the heavier truck weight. Wider roads in a pit will necessitate a higher stripping ratio, since the pit walls will be pushed back to the waste area, and this will increase the mine development costs.

Consider this figure in which we have an orebody shown in blue and two different pit limits for two different roadway widths. The blue pit limit corresponds to a wider road than the orange pit limit. As you can see in the below figure, the blue pit needs additional waste removal equal to the dashed area, compared to the orange pit. Therefore, more waste is now mined out per ton of ore, and the stripping ratio will be higher.

Figure 9.1.2: Orebody with two pit limits for two roadway widths

Source: A. Lashgari. © Penn State University, is licensed under CC BY-NC-SA 4.0 [1]. 

Loss of Selectivity

We talked earlier in the course about selectivity, and you will recall that selectivity is a measure of how effectively we can extract the ore of interest while taking as little waste material as possible. Larger mining trucks need to be matched with larger loading equipment; and larger loading machines have larger buckets. A large bucket reduces selectivity in loading, and that means you will dilute more of the ore with waste material than with a smaller piece of loading equipment.

This effect can be illustrated by considering the block of ore shown here, in Figure 9.1.3(a), assume that we have two options to load the material: a large bucket shown by the blue outline in Figure b and a small bucket the size of the blue outline in Figure c. With the large bucket, we will load all of the ore in the block, shown in orange, as well as the waste material shown in white. With the small bucket, we can be more selective, and taking three small bucket loads we get all of the ore with far less waste. As you can see, the option with the small bucket will reduce the dilution, and this in turn will reduce the unit cost of production.

Figure 9.1.3  The effect of bucket size on selectivity and dilution for one block of ore.

Source: A. Lashgari. © Penn State University, is licensed under CC BY-NC-SA 4.0 [1]. 

Equipment Life Cycle

The service life of mining equipment is measured in tons, hours, or years, depending on the type of equipment. A haul truck might have a life of 100,000 hours of operation, which may translate into 15 years of service life at a particular mine, and maybe as much as 20 years at a different mine with a different work schedule. Given the high capital cost of mining equipment, much effort is given to match the life of equipment with logical periods of production at the mine. This is yet another reason why mine planning is so important, to ensure that maximum benefit can be obtained from every piece of equipment over the life of the mine.

It is vital for a mining operation to determine an operationally viable mining sequence for the deposit and, subsequently, a production schedule that is achievable and economically sound. The Mine Planner's objective is to schedule and plan operations to achieve maximum return (of profit) on investment, through capital investment (e.g., mobile equipment), mine design, production scheduling, and preparation of the mineral product according to specifications that are required for the market. Therefore, “quantity” and “quality” would be important “drivers” in mine production planning.

BEFORE MOVING ON TO THE NEXT LESSON, please take a little time to think about the concepts presented in this section of this lesson. These will be underlying themes in many other topics.

9.1.2: Mathieson's Planning Objectives

We’ve talked about the economic foundation of mine planning, and we’ve looked at several examples of how the mine design and the mine schedule can affect the economics. This is a good foundation for understanding the material in this course, as well as preparing you for more detailed studies in later courses. I want to conclude this lesson on mine planning by presenting a list of objectives to guide your mine planning decisions.

The starting point for your mine planning activities will be the geological model of your orebody; and this will be true, regardless of whether you have come into the planning process at the startup of the new mine, or after the mine has been in operation for years. You will have a block model, which details the grade of the ore in discrete categories and locations. The size of the blocks and the number of categories is a separate decision, but one that does not need to be addressed here. Consider this block model, which uses one of five grade ranges for each block. So, this is your starting point – where do you go from here? Undoubtedly, your approach will be multifaceted. There is a set of planning objectives first compiled by Mathieson and published in 1982. These objectives apply more or less to underground as well as surface mining, although his famous paper was directed at the surface mining community.

Figure 9.1.4: Block model

Source: MinSoft Computing

Here is a list of the major objectives of any surface mine planning exercise at the feasibility level:

Mining the “Next Best” ore

Mine orebody in such a way that for each year the production cost to produce a given amount of final product is minimized. What do we mean with the “Next Best”? The next best ore block here is the block that (i) maximizes the profit and (ii) does not contradict any other constraint or limitation, e.g., equipment operating room and slope design. To find the “next best” in the above figure, we will look for the red blocks that are close to the surface. We chase the red blocks because they contain a higher grade of ore that is easiest to mine. What do I mean by easiest to mine? I mean minimizing the production costs, and this normally suggests closest to the surface without any complicating factors. A complicating factor would be an environmental constraint or a geologic anomaly, for example.

Maintain sufficient ore exposure

Give yourself options on which mining face or faces will be scheduled for production on a given shift. If you normally have five active faces on any given shift, then you will want to have at least ten faces available for work on any given shift. Then, if there are unexpected ground control or other problems, you can simply move active work to a different face. Moreover, mining operations try to maintain enough ore exposure to meet production needs for an extended period, e.g., six months. This helps the company to reduce the risk that unforeseen events will adversely impact production. This is particularly true in the early years that are so critical to economic success. Unforeseen events could include a shortage in the equipment fleet due to delivery delays or significant breakdowns, labor issues due to union strikes, or difficulty in recruiting and retaining certain positions, adverse weather events, slope stability issues, and so on. Don’t forget to account for geological and engineering miscalculations. In the early stages of mine planning, geologic and operational information is limited, and this creates a greater level of uncertainty.

Maintain proper operational parameters

Adequate bench width and properly designed haul roads are essential to achieving the design level of production. Limited operating space on the bench can increase the cycle time of the operation, which will decrease productivity and production, and it may pose a safety hazard for the equipment and workers. Improperly designed haul roads can: increase the round-trip cycle time, which will affect productivity and production; greatly increase tire wear, and tire replacement represents a major cost element in many surface operations; and create safety hazards. 

Let’s take a look at this picture of an open-pit copper mine. A loading and hauling operation is identified with arrow #1. If there were not enough operating room for the shovel, it would take longer for the shovel to dig, load, and dump the material. Concurrently, the truck driver will likely require more time to maneuver into position for loading. Arrow #2 points to a loaded truck traveling uphill to the dumping point and an empty truck making the return trip down-hill to the loading point. If proper operating room is not maintained, the trucks will need to reduce speed when passing. More importantly, there may be an increased risk for a truck to lose control and then go over the edge or crash into another truck. Therefore, it is incumbent on the mining engineer to design properly the benches and haul roads to ensure that safety and production are not compromised.

Figure 9.1.5: Open-pit copper mine

Source: A. Lashgari. © Penn State University, is licensed under CC BY-NC-SA 4.0 [1]. 

Defer stripping

As we discussed before, we would like to expedite profit making in the early years of operation, and to the extent that it is logical and achievable, we want to minimize the amount of stripping required to access the orebody. In the process of deferring stripping, we don’t want to compromise other objectives, such as the two that we just discussed.

Minimize the risk of delays in the initial cash flow

The production rate that is used for the early years of the mine is one of the more critical variables in determining the initial cash flow. Yet, it is also the variable that is subject to significant change over which you may have little control. Equipment delivery, outside contractor performance, workforce startup and training issues, and initial geologic uncertainty are examples of factors that can halve your production target in a given year. Imagine if your financial projections as supported by mine planning missed their target by 50%! Once again, you’d be looking for a new job! Consequently, the first year or two of production is de-rated to account for these uncertainties. If you were designing the mine to produce 6 million tons per year, it would not be unreasonable to choose a target for two million for the first year and three million for the second year. These aren’t arbitrarily chosen but will be based on simulations and risk analysis. The point here is that it is important to determine achievable production levels in the first few years of the mine life.

Maximize pit slope, while minimizing the risk of slope failure

We’ve looked at a few examples to illustrate the effect of the ultimate pit slope on stripping ratios and economics; and the “take-away message” is that the steeper the slope, the more favorable the economics! Unless, of course, you have slope failure! And then, once again, you are looking for a new job! Just kidding ... Even with very good ground control, slope failure can occur. It’s like walking a tight rope. The slope stability calculations are limited by geologic uncertainty, and so, safety factors are required. But how much of a safety factor: 1.5, 2.0, 4.0, or 10? There are analytical methods in rock and soil mechanics to improve the accuracy of your predictions, and the use of risk analysis is standard. Nonetheless, the possibility of a slope failure weighs on the operations and engineering personnel. A clever and necessary approach is to choose a smaller safety factor, but utilize sophisticated slope monitoring instrumentation and techniques to detect the earliest signs of incipient slope failure. Engineering and operational interventions can be initiated immediately to prevent a catastrophic loss.

Optimize cutoff grade and production rate

Mining companies perform market surveys to make an accurate estimation of the amount of product that can be sold every year, and at what price. Remember, the cutoff grade was based on an assumed selling price and an assumed mining cost. As the market changes, the cutoff grade will change, and as production costs change, for better or worse, the cutoff grade will change.

As an example, the annual tonnage that can be sold is suggested by the marketing surveys for a copper mining operation. The annual production rates for the mill and mine are calculated accordingly. Keep in mind that mill production rate only depends on ore, whereas the mine production rate includes both ore and overburden removal. The mining equipment should be selected in such a way that sufficient ore is sent to the mill, and so, there is sufficient capacity to maintain additional faces to satisfy the “sufficient exposure” constraint. There are flexibilities for both ore production and waste mining plans. Moreover, multiple destinations may be considered for the material, including sending material to a waste dump, a low-grade leach dump, a high-grade leach dump, the mill and sometimes an ore stockpile, which is used to feed the mill when mine production is down. Cutoff grades need to be determined for each of these destinations. I imagine that you are beginning to appreciate the need for thorough mine planning to support these decisions — hence, this objective to examine the economic merits of alternative production scenarios, including different ore production rates and cutoff grades, for the purpose of optimizing the cutoff grades and production rates.

Optimize method and equipment selection

Earlier in this module, the importance of selecting the proper size of mining equipment was discussed. We learned that all of the possible options for the fleet should be studied and the best option should be selected. Mining equipment selection is a complex multi-criteria decision-making problem. These parameters include, but are not limited to, the unit cost of the operation, equipment availability, selectivity, operating space, environmental impact, and many other technical and economic parameters. A mine planner should subject the proposed mining strategy, mine development plans, equipment selection, and environmental protection plans to very thorough "what if" contingency planning.

This is also a good time to remind you that mining methods and equipment can evolve and change over the life of a mine. Just because you started out with a specific method and unit operations, does not mean that you have to stick with those for the next 30 years! You may want to change one or both based on new conditions, new technology, and so on. This can only be assessed through ongoing engineering simulations and analyses.

This concludes the summary of mine planning objectives as first outlined by Mathieson. While these eight objectives are timeless in their guidance, i.e., they were true 30 years ago and will be every bit as true 30 years from today, we need to add one additional objective to reflect society’s nascent and evolving value of sustainability.

Develop sustainable mine plans to achieve sustainable mining operations

We talked about sustainability in mining and its importance. If we are to mine in a sustainable manner, then our mine planning must reflect that value. This means taking steps to avoid sterilizing the reserve, maximizing recovery rates, minimizing environmental impact, and ensuring worker safety and health through the design of the mine, the selection of equipment, and the choice of unit and auxiliary operations. Much, but not all of the mining industry has been the vanguard of the triple-bottom line: economic, safe, and environmentally responsible operations. The triple bottom line concept has evolved and changed into the term sustainable operations. It will be your responsibility through mine planning to ensure that the industry continues its journey to sustainable mining in all commodities and in all locations.

Now, armed with an understanding of mine planning, let’s take a look at surface mining methods!

9.2.1: Open Pit Mining

Open pit mining almost always applies to non-coal materials, mostly metal and aggregate mining. However, near-surface steeply dipping coal seams are extracted using open pit mining. Reclamation, i.e., returning the land to an acceptable standard of productive use of the open pit mine is deferred until the mine closes. Let’s look at a few diagrams to better understand the relationship between the suitability of open pit mining and the spatial characteristics of the deposit.

Figure 9.2.2: L: steeply pitching deposit, and R: massive deposit

Source: A. Lashgari. © Penn State University, is licensed under CC BY-NC-SA 4.0 [1]. 

On the right side, you can see a massive deposit, and on the left, there is a steeply pitching deposit.

Now, try to imagine the way that mining should proceed to extract the ore in each case. You may think of the “open pit” as a huge upside down cone that is superimposed onto each deposit. In the case of the deposit on the right side, we would have this.

Figure 9.2.3: Massive deposit with cone representing open pit

Source: A. Lashgari. © Penn State University, is licensed under CC BY-NC-SA 4.0 [1]. 

In the case of the steeply dipping deposit on the left, our open pit would look more like this.

Figure 9.2.4: Steeply pitching deposit with cone representing open pit

Source: A. Lashgari. © Penn State University, is licensed under CC BY-NC-SA 4.0 [1]. 

These diagrams also provide insight on the reason that reclamation has to be delayed until the end of the mine life. Simply put: we start mining from the surface and work our way, level by level, down into the deposit. If we were to return the waste material to the place where we mined it, we would block access to the ore in the lower levels.

Did you notice in the diagram of the steeply pitching deposit, the pit bottom stopped well before reaching the end of the orebody? As we go deeper, our stripping ratio will eventually become prohibitively large. Moreover, as we go deeper, mining costs, other than stripping, will increase also. For example, the time and cost to transport ore from the lower levels of the pit to the processing plant will increase to the point at which we will have to buy additional trucks to maintain production levels. When we reach the point where the cost to recover the ore exceeds its value, we have two major choices: begin the mine closure stage of the mine’s life; or begin an underground mining operation. The latter is common in metalliferous deposits. Of course, that point may not be reached for decades, as the depth of the pit may exceed 2000’ in some cases.

In the previous diagrams, I illustrated the sides of the pit with a straight line. Open pits cannot be operated in this way. Instead, the pit walls need to step down, which provides access to the pit by people and equipment, and it provides a working platform at each level. We refer to these steps as benches, and this diagram is a more accurate representation of the pit than are the previous ones.

Figure 9.2.5: Diagram to show an open pit with benches.

Source: A. Lashgari. © Penn State University, is licensed under CC BY-NC-SA 4.0 [1]. 

The number of benches will depend on the bench height and the ultimate pit depth. In the last lesson, the concept of slope stability was introduced, and that will weigh heavily in the decision of the number of benches, their dimensions, and the angle of the pit wall. It is incumbent on the mining engineers to ensure sure that slope instabilities will not occur throughout the mine life, as we discussed in the last lesson.

9.2.2: The World's Largest Single Pit by Size

Here is an aerial view of one of the largest and most famous open pit copper mines – the Bingham Canyon Mine operated by Rio Tinto.

Figure 9.2.6: Bingham Canyon open pit copper mine.

Source: Mining Global [66]

Rio Tinto Kennecott runs the world’s largest single pit by size, Bingham Canyon open pit copper mine. This pit stretches three-quarters of a mile deep and nearly three miles across at the top. The pit is wide enough to stack 12 aircraft careers end-to-end. This mine carries out a sequence of drilling, blasting, loading, hauling, crushing and conveying 24 hours a day, seven days a week, 52 weeks a year. It has been in production for more than a century, and has produced more than 19 million tons of copper. Gold, silver, molybdenum and sulfuric acid are byproducts of Bingham Canyon Mine. The mine has 12 shovels (10 electric, two hydraulic) that fill an average of 2,100 haul-truck loads each day. There are 110 haul trucks ranging in capacity from 240 to 320-tons. You might not expect this, but there are seven water trucks with 50,000-gallon tanks, and these trucks spray water onto haul roads round-the-clock to reduce dust. Ore is transported to, and unloaded into, one of the world’s largest crushers, running at a rate of 10,000 tons per hour. At the end of this lesson, I am going to direct you to a video where you will learn more about these large open pit mines.

9.2.3: Pit Terminology

Now, let’s learn the terminology associated with the pits. Using this figure, we can define key elements.

Figure 9.2.7: Key elements of an open pit mine.

Source: A. Lashgari. © Penn State University, is licensed under CC BY-NC-SA 4.0 [1]. 

• Bench floor: Horizontal surface of the bench
• Bench face: Vertical surface of the bench
• Face angle (α): The average angle the face makes with the horizontal
• Bench height (H): Vertical distance between upper and lower bench surface
• Toe: Interior vertex formed by the intersection of the bench face and bench floor
• Crest: Exterior vertex formed by the intersection of the bench floor and bench face
• Bench width: The distance between the crest and the toe measured along the upper surface
• Bank width: The horizontal projection of the bench face

There are two types of benches in each open pit mine: working benches and safety (catch) benches.

Working Benches

Working benches are where unit operations (drilling, blasting, loading, and hauling) are performed. Working benches are wider than safety benches to provide enough room for the mining equipment to perform their tasks.

Safety (Catch) Benches

Safety benches are designed to collect the sliding material from the benches above and stop the downward progress of large rock pieces or boulders.

Bench height

Bench height is one of the first parameters to be determined during the mine design process. Can you guess the parameters that should be taken into account when determining proper bench height?

Bench height depends on several parameters:

Overburden and Ore Properties

We discussed before that ultimate pit slope depends on material (ore and waste) properties. These properties include rock strength and the number and direction of rock discontinuities. Benches are just like smaller pit walls with a much lower height (usually 20 to 50 feet). Therefore, benches should also be designed in such a way that no risk of instability is created due to excessive height of benches. There is a direct relationship between rock strength and safe values for the bench face angle and bench height.

Production Rate

Production rate of a mine is a parameter that is usually determined by market demand and selling capacity. When the required production rate is determined based on the market situation, then the annual mine production rate that includes ore and waste extraction is determined in such a way that enough product for the sale purposes is guaranteed. Higher benches are normally required to achieve higher production rates.

Degree of Selectivity

We discussed selectivity in the last lesson, and the definition of the parameter is unchanged. However, insofar as bench height is concerned, there is a selectivity issue, which is illustrated with these three figures, below. These figures (are attempting to) illustrate local variations in grade, and the effect of different bench sizes on the selectivity.

Figure 9.2.8: Local variation in grade and the effect of different bench sizes on the selectivity.

Source: A. Lashgari. © Penn State University, is licensed under CC BY-NC-SA 4.0 [1]. 

As you can see, the smaller bench height results in better selectivity. Better selectivity means less waste is mined, loaded, hauled, and run through the mineral processing plant; and this translates into lowered mining and processing costs. These mineral beneficiation plants are designed to work best when the characteristics of the feedstock into the plant lie within a specified range. If your feed into the plant is outside of the range, the plant’s efficiency will suffer. You may not recover all of the mineral, for example. Operating mines will use a blending strategy, where on a truck-by-truck basis, they will blend higher-grade and lower-grade ores to achieve a feed that is within the specified range. The more selectivity you have, the more effectively you can blend.

Size and Type of Equipment to Meet the Production Requirements

After the annual production rate of the mine is determined, the required daily and even hourly production rates will be calculated based on assumptions about the number of working days per year and the number of shifts per day. Then mining equipment is selected so that these production requirements can be met. Loading equipment is usually the first to be selected; next, appropriately sized trucks are chosen to match with the loader, and then the complete fleet of loading and hauling equipment will be specified. The maximum reach of the loading equipment should be roughly the same as the bench height.

Regulatory Limits

Federal and state regulatory agencies have regulations that limit bench height for specific mining operations, depending on the commodity, to ensure the safety of mine workers and equipment inside the pit. The engineering calculations may indicate that a greater bench height is safe, but it is unlikely that the regulatory agency would allow an exception to their rule.

Bench Face Angle

Bench face angle is ultimately limited by geotechnical considerations, as we generally want to have as steep of an angle as we can sustain without an unacceptable risk of a slope failure. Despite excellent engineering and slope monitoring, failures can occur. This photo was taken in 2013 shortly after the slope failure at the Bingham Canyon Mine. Look at the shops and warehouses in the picture to appreciate the size of this failure. It is truly a miracle that no one was killed – the potential was there for massive loss of equipment and multiple fatalities.

Figure 9.2.9: Slope failure at Bingham Canyon Mine

Source: Mining Global [66]

Please watch the following video (20:11). It will give you a good overview of the open pit mining method, as well as a bit more background on the slope failure. The first half of the video presents an interesting historical perspective, and the last half shows contemporary mining practices.

Safety Berm

A safety berm is a pile of broken material constructed along the crest to form a “guard rail” to prevent trucks or other equipment from inadvertently driving over the edge of the haul road. It also serves to catch falling rocks. The height of the safety berm should be greater than or equal to the tire radius of the largest truck. These berms are required by federal regulation, but even with berms, accidents can happen. The picture on the left shows a haul truck, or what remains of it, after going over the highwall; and the one on the right shows a piece of rock that slid off the highwall and landed on a truck below. The former most certainly resulted in a fatality, whereas the latter probably did not, in this instance.

Figure 9.2.10: L: Remains of a haul truck, and R: Rock sitting atop a haul truck.

Source: Mining Mayhem [66]

Haul Roads

These are added to the pit walls to provide access between benches and the plant next to the pit. Open pit haul roads have a gradient up to 15%. The width of the haul road is designed based on the width of the largest truck that travels in the pit and the number of the lanes that are required to provide proper access for haul trucks, auxiliary equipment such as explosives trucks and water trucks, and smaller vehicles, such as pickup trucks, while minimizing congestion.

Now, let’s review the bench geometry concepts and open pit elements by looking at two annotated photographs.

Figure 9.2.11: Diagram showing bench width and height, bench floor and bench face, and face angle.

Source: Vladislav Kecojevic & A. Lashgari. © Penn State University, is licensed under CC BY-NC-SA 4.0 [1]. 

Figure 9.2.12: Diagram of open pit mine with major features labeled.

Source: SlideShare.net  Hassan Harraz [67]

9.2.4: Mines in the U.S.

Here is a list of some prominent open pit mines in the U.S.

As you can see from this list, most of the US largest open pit mines in the U.S. are located in the state of Nevada and Arizona, where there are large deposits of gold and copper-molybdenum.

The sequence of steps for the development of an open pit mine is as follows:

• determining the three-dimensional distribution of mineralization and grade;
• establishing the economic limits for the pit;
• selecting suitable sites for waste embankments and soil stockpiles;
• clearing vegetation from the land intended as sites for pit and waste embankments;
• siting of processing, maintenance, office, and transport facilities close to the pit but outside potential pit limit expansion;
• selecting equipment; laying out haulage roads;
• initiating “pioneering” cuts in the form of “box cuts” or access roads.

Acquisition of rights, permitting, and construction of facilities are not shown separately in this list, but are an overall part of the development sequence.

Rio Tinto is a global mining company that owns the Kennecott Bingham Copper Mine, and they have produced an informative virtual tour (may be viewed at your own pace) of the mine and related operation. It is outstanding. You should watch it several times to reinforce the content of these lessons as well as to see vivid images of the equipment and operations in action. Follow the instructions below to get started.

• Go to the Kennecott homepage. [68]
• Scroll down until you see the virtual tour of our operation.
• You will be asked to provide your name and email address. You may do so if you choose, but if you don't want to do so, simply click the "X" in the upper right corner of the dialog box.

9.2.5: Stone for Construction, Chemical, and Other Uses

Before leaving this lesson on open pit mining, I want to mention an important application of this method – stone for construction, chemical, and other uses. You’ll recall from our initial discussion on mining methods that I told you that the common term for this mining is quarrying, not open pit mining. As mining engineers, that does not thrill us because quarrying is the correct term for the mining of dimension stone. However, it doesn’t much matter what we think, because millions of people use the word quarry and quarrying to refer to these open pit operations where stone is mined. Frankly, I suggest that you surrender, and refer to them as quarries as well! That’s what I am going to do here!

If you recall the graphic at the beginning of Lesson 9 [69] in this module on surface mining, these stone quarries account for more than a third of all mines in the U.S., numbering over 6,000! The majority of the stone that is mined is crushed and used for construction – roadbed, concrete, bituminous concrete, erosion control, and so on. Limestone and sandstone are typically mined, but in some areas, an igneous rock known as trap rock is mined. These construction uses are relatively low value, with the product selling for around $10 per ton. A small but important market exists for chemical-grade limestone, manufactured sand, and specialty products such as kitty litter or the small grains incorporated into the surface of roof shingles. These markets will pay upwards of hundreds of dollars per ton for the mined product. Given the importance and size of the quarry industry, I wanted to make a few additional comments on this application of open pit mining.

Everything that we have discussed under the topic of open pit mining applies to quarries, i.e., open pits in which crushed stone is mined. The planning, engineering, equipment, and so on, are the same. Most quarries are smaller than the large open pit mines for recovering metals, and as such, the equipment selection may be different. Wheeled loaders are more likely to be used than shovels, and the haul truck capacity will be smaller. Hydraulic excavators may be used, but it would be rare to find a shovel. The safety, health, and environmental challenges are similar, and the engineering design and production engineering is similar. Looking at this quarry, for example, it would be difficult to distinguish it from any other open pit mine.

Figure 9.2.13: Equipment at work in a quarry

V. Meleshko/Shutterstock.com, used with permission

If you take MNG 441, you will learn more about the design of these open pits. Also, we offer an elective course focusing on aggregates, with an emphasis on stone quarries.

9.3.1: Area Mining

In the left figure, we have relatively constant thicknesses for both overburden and ore. Our first step is to remove the surface vegetation and stockpile the topsoil for later use in reclamation. We will only remove the vegetation and topsoil in the area that we will mine within a period of weeks. Next, we need to gain our initial access to the seam, and we will do this with a box cut. Depending on the hardness and tightness of the overburden, we may need to drill and blast to facilitate its removal. If it is soft enough, we’ll remove it without the need to drill and blast. Depending on the thickness of the overburden, and the required size of the box cut, we will probably use a dragline to make the box cut, although it could be done in certain cases with dozers and scrapers. The overburden from the box cut will be placed to the side of the pit. When the coal seam under the box cut is exposed, we start mining the coal. After the coal in the box cut is removed, a subsequent cut is accomplished and the waste material from the second cut is placed in the space created by the initial box cut. This process continues by making subsequent cuts, putting extracted overburden in the former cut and mining the uncovered coal seam. The collection of sequential cuts forms a strip, and the strips will traverse a portion of the property. Take a look at this figure, and don’t be concerned about the difference between the low wall and the highwall. You can see the box cut. In the lower part of the figure, you can see the strips being mined and the overburden or spoil placed into the adjacent and previously mined strip. The overburden swells in volume when it is extracted, and that’s why you see the piles of spoil protruding above the original contour of the land.

Figure 9.3.2: Diagrams to show a box cut, strip cuts and spoil piles

Source: Humphrey, 1984

Sometimes, it is difficult to imagine the cuts and strips and how it all progresses. As you look at the pictures and diagrams in this lesson, it will come together for you. Another visual for right now would be to think of a bowling alley with its many bowling lanes. You can think of the bowling alley as the area to be mined, and each lane as a strip. Within each strip, a series of cuts will be made so that the “mining activity” can progress from the beginning to the end of the strip. Mining of the first strip is a little tricky, but once it has been completed then all of the other strips will be completed as shown in the diagram above – remove the overburden, expose the coal seam, mine the coal, cast the overburden from the next cut into the previously mine strip.

Overburden removal in US area mining is commonly accomplished with large draglines, stripping shovels, or in small mines, by conventional excavation and haulage (truck-shovel or dozer-scraper) methods. In some other countries such as Germany, bucket wheel excavators are also used to remove overburden. Dozers and scrapers are also used in clearing vegetation, removing topsoil and subsoil, establishing level benches and surfaces for the dragline, as well as many other activities that involve pushing the spoil. In the next lesson, we’ll look at a video of area mining to recover phosphate, and you will see the prep work done by the dozers as well as the operation of the dragline.

Here is a picture of a Black Thunder Mine. You can see the individual strips where overburden has been placed; and within each strip, note the individual piles. Each pile corresponds to a cut. The uncovered, but unmined coal seam is visible as well. Given the thickness of this seam, on the order of 100’, a large shovel will be used to mine it. If it were thinner, say 10’, then a small shovel could be used, but more likely a wheeled loader. Often, the coal seam is drilled and blasted to facilitate rapid loading of the coal.

Figure 9.3.3: Black Thunder Mine with labeled coal seam, dragline, spoil piles, and blasted overburden

Source: High Country News [70]

There are a few other elements within this picture that I’d like you to note. Do you see the blasted overburden, and where is the dragline sitting with respect to the top surface of the blasted overburden? Quite a bit lower, right? So, how does this happen?

A dragline has a limited reach, and there is a tradeoff between its reach and the size of the bucket. For example, it may have a reach of 200’ with a 200 yd³ bucket, and 300’ with a 120 yd³ bucket. Without getting into the design details, optimizing the reach and bucket size combination sometimes requires that the overburden be removed in two separate passes. The interesting thing about the process in the Powder River Basin is that they remove up to 40% of the overburden by a process known as cast blasting. In cast blasting, not only does the blast loosen the overburden, but the round is designed such that up to 40% of the overburden is actually thrown or cast into the adjacent strip! They may use a million pounds of ANFO or more to accomplish this!

Dozers are then used to push more of the overburden into the adjacent cut, and importantly, to create a flat and level surface for the dragline to operate. You can also see a temporary haul road developed by the dozers to provide access to the coal seam for haul trucks and to the dragline pad for smaller vehicles to service and supply the dragline.

I want to point out one last detail before leaving this picture. Did you notice the shape of the pile of overburden? They are all very distinct and nearly identical in shape. Do you think that is a coincidence? The piles of material form at an angle known as the angle of repose. This is the angle that a dry, granular, material in a very loose condition will form with respect to the horizontal. The angle of repose is a unique characteristic of the material, whether it is iron ore, coal, #2B crushed stone, flour, or sugar. If we know this angle, we can calculate all sorts of interesting things, such as the maximum angle of a conveyor belt so that the ore will not tumble back on itself or the size of a stock pile.

When it is time to reclaim the strip, bulldozers and scrapers work together to spread these spoil piles. Then the subsoil and top soil are placed back to reclaim and return the land to its original contour lines. Often, the rock layers immediately above the coal seam will be “acid bearing”, i.e., minerals such as pyrites that will create an acidic drainage when ground water percolates through those layers. Acid mine drainage is a great environmental concern, and to reduce the level of acidity under the reclaimed lands, crushed limestone is sometimes added to neutralize acid bearing layers. This process is called liming.

Here is another picture to illustrate the operations of the open cast area mining method.

Figure 9.3.4: Labeled photo of open cast area mine showing a drill, a haul road, drill holes, and spoil piles

Source: Washington Times [71]

Draglines

Figure 9.3.5: Diagram to show three primary dragline parameters

Source: P&H Mining (Komatsu / Joy Global) [72]

Please take a minute and think about the way that each of these three parameters plays a role in the design and the width of the strips.

Next let’s take a look at the key components of a dragline. You should understand the purpose of the drag, hoist, and dump ropes. You may want to think about the challenges of the structural design of the boom itself, and the tradeoff between bucket capacity and boom length. What about moving the dragline? How does it move about the bench? Very slowly and with great difficulty!!! There are no wheels nor crawler tracks because it is so massive and so heavy that conventional approaches would fail in all but the smallest draglines. Instead, the dragline walks! It sits on a giant cast iron tub so that its enormous weight can be distributed over a large area. On either side of the tub, there is a massive shoe. The shoe is connected to an eccentric cam, and to move, the shoes lift the tub off the bench and propel it forward where it comes to rest. And then, the cycle is repeated. Fortunately, the dragline does not need to move quickly in the course of normal production. However, if you need to move the dragline some distance, say to an adjacent property located on the other side of a highway, it can be a tedious process – like watching grass grow!

Figure 9.3.6: Key components of a drag line

Source: P&H Mining (Komatsu / Joy Global) [72]

Now, let’s look at the operating cycle of the dragline in more detail, with the help of these diagrams.

Figure 9.3.7: Operating cycle of a drag line

Source: from Introductory Mining Engineering, 2nd Ed., Hartman, H.L, and J.M. Mutmansky

The dragline is positioned on top of the blasted material, which has been leveled and prepared by a dozer. The drag bucket is positioned at its furthest reach, lying on the overburden. The drag rope will begin to pull the bucket towards the tub, and as the drag bucket is pulled, it will dig into the material and the bucket will be filled. Once filled, the dragline operator will begin to hoist the bucket and will begin to swing the boom toward the spoil pile. By the time the bucket reaches the spoil pile, the operator will have completed hoisting the bucket, and will then dump the bucket. Once dumped, the operator will swing back to the position to dig, while at the same time, dropping the drag bucket so that when the swing cycle has been completed, the drag bucket will be lying in position to begin the cycle all over again. A typical time to complete one cycle is 45 seconds.

Bucket Wheel Excavators (BWEs)

Let’s close out this discussion on area mining with a brief mention of bucket wheel excavators (BWE), shown in the picture here. BWEs are massive machines to quickly remove overburden. To gain an appreciation for their size, look for the front-end loader in the lower left part of the picture. Did you find it? These machines have been used successfully in the brown coal fields of Germany for many decades. Some BWEs were imported into the U.S nearly 50 years ago. Unfortunately, the results were disappointing. The characteristics of the overburden were unsuited to the capabilities of BWEs. The bucket wheel excavator is best used for soft and loose overburdens. While these were common in the German lignite fields, they were less common here in the U.S. Consequently, they were never adopted for use in the U.S.

Figure 9.3.8: Bucket Wheel Excavators at work

Source: Wikimedia [73] 

9.3.2: Contour Mining

It would be great if all coal were found in conditions similar to those in the Powder River Basin: relatively flat and wide-open space, thick coal seams, and shallow cover (overburden). Mother Nature was not so accommodating! A lot of high quality thermal and metallurgical coal is found in rugged terrain, in areas such as the Appalachian Coal Field, which runs from Pennsylvania down to Alabama. Contour mining methods evolved to allow the economic extraction of coal using smaller and more agile equipment with more flexible mining patterns.

As with area mining, it is necessary to remove the vegetation and remove the overburden to gain access to the coal seam. There are two ways to handle the removed overburden: haul it all the way down the hill and put it into a chosen valley in horizontal layers; or to place it in the void of the cut, where the coal has been already removed. Mining will proceed by advancing into the hillside until reaching the break-even stripping ratio. At this point, the highwall will remain. Under certain conditions, we may choose to take advantage of the exposed coal seam, and develop an underground mine or practice highwall mining, which will be discussed later in this lesson.

Now, let’s take a look at a schematic view of contour mining approach and then a photo of a contour mining site. The diagram shows more than one coal seam. In fact, it is not uncommon for multiple seams to be recovered. It creates a small sequencing and scheduling challenge because you have to remove the interburden between the seams, and place it, but it is commonly done.

Figure 9.3.9: Schematic view: contour mining approach

Source: "What is Strip Mining? [74]" The Assay.

Note the highwall in the photo, and the exposed coal seam.

 

Figure 9.3.10: Contour mining site

Source: Giles [75]Asford [75]

Haulback Mining

An important submethod within contour mining is haulback mining. Haulback mining, which is illustrated nicely in this diagram from the textbook, is a method in which overburden removal and then coal extraction follows the outcrop, proceeding in a series of nearly rectangular pits. As mining is advancing at one end of the rectangular cut, reclamation is occurring on the back end. In this method, mining roughly follows the contours of the land, and the ultimate pit width is dictated by the economic stripping ratio and steepness of topography.

Haulback mining starts with developing a box cut and hauling the overburden to a suitable disposal site. Coal is mined out and mining progresses with removing the next cut of overburden, and hauling it by truck, scraper, or conveyor to fill previously mined-out cut. This method is called haulback mining, because waste moves in a direction opposite to the direction of mining, as illustrated in the diagram.

Figure 9.3.11: Diagram of contour mining site in action and employing the haulback method

Source: Mathteck Inc., 1976

Mountaintop Removal

Mountaintop removal is another submethod of contour mining, where a flat lying coal seam occurs near a mountain top and outcrops on opposite sides of a hill. When the overburden is sufficiently thin (less than 165 feet) across the hill, the deposit can be mined from outcrop to outcrop. In this method, the hill is completely removed, and the spoil is leveled rather than contoured to approximate the original topography, creating a flat land. This method was commonly used in Appalachia, but is very controversial, and is no longer recommended. The below picture shows a mountaintop removal operation in southern WV, prior to reclamation of the site.

Figure 9.3.12: Aerial photo of mountaintop removal site

Source: Ohio Valley Environmental Coalition [76]

9.3.3: Auger Mining

Earlier in this lesson, you saw the highwall that remains after the ultimate pit limit is reached in a contour mining operation. At that time, I pointed out the possibility of recovering the coal under the highwall, by either developing an underground mine or highwall mining. The former is rarely economically feasible today, but the latter is commonly practiced. A distinction is made between auger mining and highwall mining, and this will become clear shortly.

In auger mining, auger holes are bored horizontally into the coal seam from the last contour cut at the high wall. As the auger rotates and advances forward, coal is cut and pushed out of the hole by the rotating action of the auger. The action is similar to that of a carpenter boring a large hole in wood using an auger bit, and that is where the name derives. Auger mining recovers around one-third of the coal under the highwall, for whatever depth of penetration is achieved. Auger cutting heads can be as large as 7 feet in diameter and may be drilled to a depth of more than 300 feet. An auger mining machine is shown here.

Figure 9.3.13: Auger Mining Machine

Source: Newcastle Herald [77]

A highwall, showing the auger holes is shown in this picture:

Figure 9.3.14: Highwall with auger holes

Source: OilPrice.com [78]

When the ultimate pit limit is reached, it is difficult to “pack up and leave” when you see the exposed coal seam. So, in some regard, the company looks at this secondary recovery technique, i.e., auger mining, as the “icing on the cake.” It amounts to coal recovery at a very low cost. It is, however, somewhat problematic and controversial. The low recovery, less than 35%, means that the majority of the reserve remains in place, but it has been sterilized, i.e., left in a condition where it will be essentially impossible for anyone to recover the remainder in the future. The holes into the side of the hill create drainage holes for acid-laden water, and that is a big problem. Surface subsidence and spontaneous combustion in the auger holes are additional concerns.

9.3.4: Highwall Mining

Highwall mining is very similar to auger mining in its concept, where the entry into the coal seam in made at the highwall of the ultimate pit limit. The difference lies in the use of a more sophisticated machine, known as a highwall miner. The highwall miner shares some common elements with a continuous mining machine. It has bit-laced cutting drum, and the coal is fed back through a materials handling system to the bench by the highwall, where it is loaded out. That’s where the similarity stops.

The highwall miner is equipped with sophisticated monitoring and a remote control system. It is important to maintain accurate control of the cutting head so that it remains in the coal seam, rather than drifting up or down into the adjacent rock layers, and so that a constant rib thickness is maintained between holes. Unlike the continuous miner, where a shuttle car is waiting at the tail to load out the coal, the highwall miner needs to move the coal from the cutting head to outside the highwall. Essentially, there is a flexible conveyor train following the cutting head. There is a significant amount of support equipment required outside of the hole, adjacent to the highwall, to provide power and handle the cables and hoses. The computerized supervisory control system and operator station are located within the highwall miner’s structure that is located adjacent to the highwall. This picture with the artist’s cutaway shows a highwall mining machine.

Figure 9.3.15: Artist's representation of a highwall mining machine in action.

Source: PRWeb, via PRWeb limited license 

What do you gain with the added complexity and cost of a highwall miner over an auger? The highwall miner can advance 1200’ under the highwall compared to a typical 300’ for an auger. Recovery rates improve from roughly 30% for an auger to near 50% for a highwall miner. The highwall miner can adapt more easily to varying seam thickness or other changes under the highwall. Its production rate can be much higher than an auger, and, in general, the unit cost of coal mined with the highwall miner will be less than with an auger. The risk of worker injuries is less with a highwall miner, primarily because the likelihood of a cave-in is less than with an auger; and attempting to remove an auger from a caved hole is fraught with dangers. On the negative side, the highwall miner does not reduce the concern over acid mine drainage.

Here is a photo of the highwall showing the holes created by the highwall miner.

Figure 9.3.16: Holes created by the highwall miner

Source: ThingLink.com

9.4.1: Dredging or Dredge Mining

A dredge is the principal piece of equipment used in the dredging method and, essentially, a dredge is a boat containing specialized mining and materials handling components. Accordingly, a dredge requires a body of water in which to operate. In many cases, this is a natural body of water such as a river or a lake, but in others, it is a man made pond or small lake. Of course, the only reason for floating a dredge is to recover something of value at the bottom of this body of water. We need to add one more condition, and that is: the material of value on the bottom must be unconsolidated, such as sand and gravel, or it must be very soft. Simplistically, the dredge is designed to lift these materials of interest from the bottom up into the dredge. Shortly, we will look at how the payload is moved in a little more detail, but for now, let’s talk for a moment about the kinds of materials that are typically recovered with this method, dredging or dredge mining.

The bottom of rivers, lakes, and harbors is often a good source of gravel. Gravel can be used sometimes in concrete as well as for a variety of other purposes such as architectural and landscaping purposes. Dredging to remove this material has the additional benefit of deepening the channel or harbor, and sometimes that is the primary purpose of dredging, and the recovery of minerals is a secondary benefit… the “icing on the cake,” so to speak!

Glaciers once covered a significant portion of the Earth’s surface, and the movement of these glaciers created extensive unconsolidated deposits of materials containing not only sand and gravel, but gold, tin, diamonds, and other heavy minerals. These alluvial deposits, created by glaciation, are also known as placers. I mention this here because you will sometimes hear or read about placer mining or alluvial mining. Although these terms may be used interchangeably with dredging or dredge mining, you can’t assume that to be true in all cases. As we will see, hydraulic mining may be used in these deposits as well. Regardless, if we have a body of water covering a placer deposit, we will consider strongly using dredge to recover the minerals of interest.

Moreover, if we have such a deposit that is not underwater, but is in an area that could be easily flooded, we will consider making our own lake and then using a dredge to recover the bottom materials. If the deposit is in an area with a very shallow water table, we may simply have to remove several feet of overburden, e.g., vegetation and soil, and the excavated area will fill with water on its own. Then, we can float the dredge and mine the deposit. There are other circumstances where we could create a man made lake to mine the deposit, but it is a complicated process, because we cannot take any action that could have an adverse environmental impact.

Let’s talk a bit more about the dredge itself. A dredge is defined by the way in which it recovers the ore from the bottom. The four types are bucket-wheel, ladder, clamshell, and suction. The choice will depend on the depth of the deposit below the surface of the water and the degree of consolidation of the deposit. The size of the dredge will depend on the desired production rate and the characteristics of the body of water, e.g., depth and extent. Let’s start by looking at a picture of a dredge.

Figure 9.4.1: Bucket wheel dredge

Source: Alibaba [79]

This is a large dredge, and specifically, it is a bucket-wheel dredge. You can see the bucket wheel on the left side of the picture. The bucket wheel rotates, digging into the soft material at the bottom. The dug or "mined" material is then transfer onto the dredge. Typically, some sort of a gravity separation is employed to segregate the material of interest from the silt, mud, and other detritus of no interest. The latter is then immediately returned to the water.

The bucket-ladder dredge is probably the most common type of dredge, as it is the most flexible method for dredging under varying conditions. The excavation equipment consists of an endless chain of open buckets that travel around a truss or “ladder.” The lower end of the ladder rests on the mine face—that is, the bottom of the water where excavation takes place—and the top end is located near the center of the dredge. The chain of buckets passes around the upper end of the ladder at a drive sprocket and loops downward to an idler sprocket at the bottom. The filled buckets, supported by rollers, are pulled up the ladder and dump their load into a hopper that feeds the separation plant on the dredge. After the valuable material has been removed, the waste is dumped off the back end of the dredge. Here is a picture of a bucket-ladder dredge, which gives a clear front view of the bucket ladder.

Figure 9.4.2: Ladder dredge

Source: DredgePoint [80]

Here is a view of a bucket-ladder dredge used to mine phosphate.

Figure 9.4.3: Bucket-ladder dredge

Credit: K. Hutton, © Penn State University, is licensed under CC BY-NC-SA 4.0 [1]

And here is that same dredge in a photo taken from a distance.

Figure 9.4.4: Note the slurry pipe running to the bank

Credit: K. Hutton, © Penn State University, is licensed under CC BY-NC-SA 4.0 [1]

Finally, here a view from the operator's cab of that dredge. Notice the computer displays providing not only video images of different parts of the dredge but also sensor data that the operator can use to better control the operation of the dredge.

Figure 9.4.5: From the operator's cabin

Credit: K. Hutton, © Penn State University, is licensed under CC BY-NC-SA 4.0 [1]

The clamshell dredge, unlike the previous two, employs a batch rather than continuous process. This type of dredge utilizes a clamshell bucket that is dropped to the bottom, scoops a bucket of material, and is hoisted back to the dredge where the bucket is dumped. This dredge can operate in deeper water than other systems and handles large material, e.g., larger rocks, well. A typical cycle time would be on the order of one minute, depending on the depth of the water. You understand the drawback of batch or discontinuous systems, and consequently, this style of dredge would only be used when its unique strengths are necessary. A typical clamshell dredge is shown here. Note the ability of this style dredge to place its payload on the dredge or on a nearby barge or structure.

Figure 9.4.6: Subsea clamshell dredge

Source: IP SubSea.com [81]

The fourth type of dredge is a hydraulic dredge. Imagine a big vacuum cleaner with a long hose – the hose is dropped to the bottom, the “vacuum” is turned on, and the material is literally sucked up the hose and captured on the dredge. Basic physics limits the amount of “lift” that can be achieved. However, the amount of lift can be supplemented with a high-pressure spray around the suction nozzle – essentially a push-pull system. This is known as hydrojet assistance. This style of dredge is suited to digging relatively small-sized and loose material such as sand and gravel, marine shell deposits, mill tailings, and unconsolidated overburden. Hydraulic dredging has also been applied to the mining of deposits containing diamonds, tin, tungsten, niobium-tantalum, titanium, and monazite. This figure diagrammatically illustrates the use of a suction dredge. Note that the use of a hydraulic pipeline to move material off the dredge is often associated with the use of this type of dredge. 

Figure 9.4.7: Diagram of a suction dredge

Source: Dariusz Kowalczyk [82], CC BY-SA 4.0 [83], Link [84]

This picture shows a very small suction dredge, which might be used to clear a tailings pond, for example.

Figure 9.4.8: Small suction dredge

Source: Doctor Dredge, LLC [85]

9.4.2: Hydraulicking or Hydraulic Mining

Hydraulic mining uses high-pressure water cannons, known as monitors, to dislodge relatively unconsolidated material. One of the earliest applications of hydraulic mining was to break down banks of alluvium containing gold and silver. These alluvial deposits are firm, but break down quickly upon the application of the water cannons. Of course, if digging the banks of alluvium were the goal, we could use wheeled front loaders or other traditional digging and loading equipment. The goal is not only to excavate the banks but to separate the gold, silver, or other metals from the sediments in these alluvial deposits. Toward that end, we use a gravity separation with water in a device known as a sluice. The sluice, which is usually constructed, consists of a trough on a slight incline. The trough bottom may have embedded ridges. The idea is that the material laden water will flow gently down the sluice and denser materials will settle to the bottom. Toward this end, the material loosened by the water cannon is channeled to the sluice; and as it passes through the sluice, the heavier materials, e.g., grains of gold and silver, settle on the bottom of the sluice while the water and sediment are carried away.

Another important, but limited, application of hydraulicking is to break down mineral ores prior to slurry transport. Slurry transport is an important materials handling method in which a water-solid mixture is pumped over some distance. In some cases, the slurry may be pumped for 20 or 30 miles, while in others, it may be hundreds of feet. One such application is in phosphate mining. The overburden is removed, typically with a dragline, and then a monitor is used to break down the phosphate matrix into a slurry, which is then pumped to the mineral processing plant. We’re going to examine a case study in the next Lesson, and at that mine, they use a dragline to excavate the phosphate matrix and place it into a small pit where it is broken down and fed into the slurry pipeline. You’ll see some good pictures of hydraulic monitors in that case study.

9.4.3: Solution Mining

Solution mining is the extraction of the ore through a dissolution process, i.e., we pump a solvent through the ore, the solvent dissolves the constituent of interest, and then we recover and process the resulting liquid, which is known as the solute or pregnant liquor. We can apply this method to in-situ deposits using a borehole mining method, or to mined material using a heap leaching method. In-situ, which implies “in-place” “in-the-seam”, and using an in-situ method, means that we extract, i.e., dissolve, the material of interest within the deposit and leave behind the gangue or waste minerals. In-situ recovery has the advantage of not disturbing the surface with a mining activity and of eliminating surface piles of processing plant tailings or waste. It has the disadvantage in some applications of creating a potentially hazardous environmental situation if the solvent or solute can migrate into the environment. Let’s talk about the in-situ methods first.

9.4.4: Borehole Mining Method

One hole is drilled into the deposit for the purpose of pumping the solvent into the deposit, and one is drilled at some distance from the other to recover the solute. The solvent flows between the two holes, creating a cavern where the ore was dissolved. For some minerals, such as common salt (NaCl) or trona, water is the solvent. For other minerals, a chemical solvent is required. Interestingly, sometimes these caverns are of greater value than the extracted mineral! In the Louisiana salt domes, for example, the caverns are used to store crude oil and natural gas. Often, multiple holes are drilled within the area of interest. Occasionally, only one hole is drilled, but the borehole is divided into an inner and outer annulus so that the solvent is forced down one, and the solute is retrieved through the other. This is the case for the Frasch process for recovering sulfur. The Frasch sulfur process uses steam to melt and dissolve the sulfur in the water, which is then forced back up the borehole; or, in other cases, the melted sulfur is allowed to drain and collected in subterranean pools where it is then pumped to the surface.

9.4.5: Heap Leaching Method

Heap leaching was first used to extract very low grades of metal ore from piles of plant tailings or waste rock. In the case of copper, for example, a solution of sulfuric acid was released at the top of the pile and allowed to percolate down through the pile. The effluent was collected and washed over scrap tin. The copper would precipitate out on the tin, the tin would then be smelted, and the copper would be separated from the tin. All-in-all, it was a fairly clever and economic process to extract metal from material that otherwise would have been discarded as waste. Unfortunately, there was no thought given to the consequences of the effluent seeping into the ground, and in the end, this practice resulted in massive pollution around the mine site. The old Anaconda Copper Company mine in Butte, Montana, for example, became an EPA Superfund Site! Today, heap leaching is practiced, but with proper engineering controls in place to ensure that it proceeds with no damage to the environment. Among other precautions, extensive containment barriers are constructed with multiple levels of redundancy.

Today, low-grade ores are “mined” on the surface and placed on these containment pads so that the valuable metal can be leached out of these “heaps.” This is an important method for recovering gold from low-grade ore that could not be economically processed by any other means. The mining of the low-grade ore is usually by large dozers, front loaders, and trucks. Usually, the depth of the excavation is quite limited, and two of the reasons for this are that as the deposit depth increases, the grade may improve to allow economic exploitation by another mining method; and the cost of excavating this low-grade ore increases with the depth, and quickly it becomes uneconomic. The time required for the leaching to be completed can vary from weeks to years, with the latter being more common; and during this period, bacterial action, as from metal oxidizing bacteria for example, is critical to the success of the process. A variation of heap leaching involves constructing the “heap” in a dump, and calling it dump leaching. Large containment vessels are being used on a limited basis as well.

Both heap leaching and borehole mining have become the favored exploitation methods for recovering uranium. This figure illustrates a heap leaching operation.

Figure 9.4.9: A heap leaching operation

Source: U.S. Nuclear Regulatory Commission [86] CC0 [87]

This figure illustrates the process at Uranerz’s Nichols Ranch Mine. This is a good example of how a borehole mining method is employed for in-situ recovery of a mineral.

Figure 9.4.10: Brochure material explaining in-situ recovery

Source: Energy Fuels [88]

9.5.1: Overview of the Phosphate Industry

Phosphate is a key component in agriculture (food production)

• granular mono ammonium and diammonium phosphate pellets for fertilizer
• animal feed supplements
• elemental phosphorous for food-additive applications.

A limited amount of elemental phosphorus is used in other industrial applications.

In 2015, Phosphate rock was mined by 5 companies at 11 mines:

• in 4 states
  • 80% in Florida and North Carolina & 20% in Idaho and Utah
• to produce 27 million tons of marketable product with a value of $2.4 billion fob mine

World Mine Production and Reserves (USGS Minerals Commodity Survey)

There are three sources of Phosphate Rock

• Sedimentary deposits of marine origin
  • Africa, China, Middle East, and United States
• Igneous and metamorphic deposits
  • Brazil, Canada, Finland, Russia, and South Africa
• Biogenic deposits from bird guano

High-quality sedimentary phosphate deposits are mined in the U.S.

The mined phosphate rock is beneficiated by a process in which the phosphate rock is reacted with sulfuric acid to produce an intermediate feedstock of phosphoric acid.

Gangue minerals associated with the deposit impact the cost of processing the ore

The gangue minerals can require larger quantities of reactants (sulfuric acid), which significantly increases production costs, and can increase time require for downstream processing, e.g., filtration; and they can interfere with the ore recovery and concentration in downstream operations, e.g., float processes.

Carbonates are more problematic. Dolomite, for example, shares mineralogical similarities with phosphate, and this makes it more challenging and expensive to remove it.

• Major gangue minerals of primary interest in U.S. operations:
  • dolomite, calcite, silica, and clays
• Certain other gangue components may need to be removed during or after the production of the phosphoric acid:
  • cadmium, radium, uranium
• Clay slimes are removed in log washers or hydrocyclones.
• Silica is removed using the well-established Crago double float process.

9.5.2: MOSAIC Florida Phosphate Operations

I suggest that you go to Mosaic’s website [89] to learn more about the company. Doing so, you will arrive at this "Who We Are" page, and you will note that Mosaic is not only a major producer of phosphate but also of potash!

Figure 9.5.1: Mosaic Company website

Source: The Mosaic Company [90] 

Their Florida operations are located in five counties of southwest Florida, as shown on this map. This is a rural area of Florida.

Figure 9.5.2: Mosaic operations in Florida counties

Source: The Mosaic Company  [90]

Resource

The phosphate rock deposit consists of a matrix of phosphate pebbles, clays, and sand, along with other gangue such as silica. (Carbonates occur in increasing amounts in the lower quality deposits.)

Key Characteristics

• The matrix is relatively unconsolidated and is typically on the order of 10’ thick.
• It is covered by an average overburden thickness of 30’ ranging from 15’ to 50’.
• The overburden is unconsolidated sand and organic matter.
• The matrix is underlain by a hard limestone.
• The water table is near the surface throughout the region.

Given these key characteristics, you should be able to choose the mining method. Can you do so? Well, at the beginning of the lesson, I told you that they employed the open cast method. Nonetheless, I would hope that you could come to that conclusion on your own. The use of hydraulicking, on the other hand, would not be apparent to you.

Let me tell you the overall sequence of operations to remove any mystery!

• Prepare the site for mining
  • dozers
• Install utilities
  • electric power
  • water drainage
  • “Pitsrdquo; for breaking down the matrix with monitors
  • slurry pumps and pipelines
• Remove overburden
  • dragline
• Mine the ore
  • dragline
• Prepare for the ore for hydraulic transport to the plant
  • hydraulic monitoring to break-up matrix
• Backfill the cut
  • dragline
• Reclaim the site
  • sand tailings pumped to the site
  • level spoil piles and grade
    • dozers
    • loaders
    • hydraulic excavators
  • add top layers
  • revegetate

We’ll take a look at many of these in more detail; but now, let’s start with an aerial view of the mine.

9.5.3: Welcome to the Operation

This picture was taken by a Penn State mining engineering intern while riding in the company helicopter!

Figure 9.5.3: Aerial view of Mosaic's phosphate operation

Credit: K. Hutton, © Penn State University, is licensed under CC BY-NC-SA 4.0 [1]

The features in this photo will become known to you as we continue.

Best Management Practices (BMP) features and land development

Off to the left side of the photo, you see a waterway that looks like a nicely manicured stream or small river. It is man made, and it is done to control the water table around the mine and ensure that the water table outside of the mine’s perimeter is not lowered or otherwise harmed. At these mines, this “ditch” is known as a best management practice ditch, because it is part of their stormwater pollution prevention plan (SWPPP) required under the Clean Water Act, and is considered a “best practice” to accomplish the intended goal.

Figure 9.5.4: The perimeter ditch
The construction of the BMPs precedes other development activities. Once completed, roadways are constructed into the site, vegetation is cleared, and the topsoil layers are carefully placed away from the active site for use during reclamation. Then, the site will be graded and leveled to provide a working surface for the draglines.

Source: J. Kohler, © Penn State University, is licensed under CC BY-NC-SA 4.0 [1]

Figure 9.5.5: Land development

Credit: K. Hutton, © Penn State University, is licensed under CC BY-NC-SA 4.0 [1]

9.5.4: Infrastructure Installation

Once a level site has been prepared, the work of placing utilities will begin. In this photo, they are placing high voltage cable (yellow in color), which will supply power for the dragline as well as the many pumps used on the site.

Figure 9.5.6: Placing high voltage cable

Source: K. Hutton, © Penn State University, is licensed under CC BY-NC-SA 4.0 [1]

As work on the electrical power infrastructure is continuing, a series of dewatering holes will be developed. This photo shows the electrical infrastructure, and do note the overhead electrical lines in the background and the electrical installation in the foreground. If you look closely, you can see 6” diameter PVC pipes protruding at the surface of the ground. These are dewatering holes.

Figure 9.5.7: Dewatering holes

Source:J Kohler, © Penn State University, is licensed under CC BY-NC-SA 4.0 [1]

This next photo gives us a close-up view, and we can see other important features. There is a small pump at the bottom of this hole, and you can see the small discharge line, where the pumped water enters a larger diameter collection pipe that is serving the entire row of dewatering holes. You can also see the electrical cable that goes down the hole to the pump at the bottom. These holes go to the bottom of the matrix, and so, a typical hole would be 50 – 60’ deep. Those pretty pink flags that you see are markers so that equipment, such as tractors and pickup trucks, don’t accidentally run over and damage pipes or electrical cables.

Figure 9.5.8: Close-up view of a dewatering hole

Source:J Kohler, © Penn State University, is licensed under CC BY-NC-SA 4.0 [1]

This next picture focuses on an electrical installation. As is often the case, this equipment is mounted on a skid so that it can be attached by a wire rope to a dozer and pulled to a new location. This equipment provides the hardware to allow a connection to overhead power lines, if required, a transformer to step the voltage down to a level required by the machines and pumps, and protective devices to allow switching and implementation of other safety features.

Figure 9.5.9: Mobile electrical installation

Source:J Kohler, © Penn State University, is licensed under CC BY-NC-SA 4.0 [1]

9.5.5: Overburden Removal

With the utilities in place, we are ready to install the pits and monitors and to begin overburden removal. Let’s talk first about overburden removal. A typical dragline in use at the Mosaic mines in Florida is shown in the next picture. Remember, we talked about how draglines move – they walk. Look at the two big walking shoes on this dragline, one on each side. The shoe is green and there is a yellow guardrail around the top surface of the shoe. As you might imagine, these things don’t break any land speed records!

Figure 9.5.10: A typical dragline

Source: K. Hutton, © Penn State University, is licensed under CC BY-NC-SA 4.0 [1]

Here is another aerial view of the mine, but in this one, you can clearly see the dragline at work.

Figure 9.5.11: Aerial view of the active mining area

Source: K. Hutton, © Penn State University, is licensed under CC BY-NC-SA 4.0 [1]

Here is a close-up of the dragline removing overburden. If you look to the left, you can see the last strip or cut that was mined (it is now filled with water).

Figure 9.5.12: Removing the overburden

Source: The Mosaic Company [90]

This short video clip (2:11) will give you a feel for the overburden removal cycle. Note the development activities ongoing to the left of the active cut.

In this video clip, you probably saw the greenish colored material at the bottom. That is the phosphate matrix, and in this next picture, you can see it as well.

Figure 9.5.13: View of the overburden and matrix shelf

Source: The Mosaic Company [90]

Evaluating the Phosphate Matrix

You’ll recall from our discussion of mine planning, that mineral processing plants are designed to perform best when the feed into the plant has characteristics that lie within a fairly narrow range. When the mined material is outside of this range, the practice is to blend loads of material from different mining faces to achieve the desired feed to the plant. This is true in phosphate mining as well. How do you know the qualities of the ore being mined at a given location on a given day? You send in the geologists to take samples and visually inspect the ore. In this next picture, you can see the geologist, accompanied by operations personnel, preparing to sample the phosphate matrix in the drag bucket.

Figure 9.5.14: Geologists sampling matrix from the drag bucket

Source: The Mosaic Company [90]

Now that we have a clearer understanding of this phase of the operation, let’s look at the dimensions of a typical strip or cut.

Figure 9.5.15: Strip or "Cut" dimensions. The mining "cut" here is 320 feet wide. The Overburden is 30 feet from top to bottom, and the Matrix is 10 feet from top to bottom. The lowest layer shown is the Hardrock Bottom, which is comprised of limestone, clay, and dolomite.

Source: The Mosaic Company [90]

9.5.6: The Mining Operation

Before the dragline can begin to mine the ore, the hydraulic monitors must be set up and ready to go. That work will be occurring while the dragline is removing overburden. Let’s start out by looking at a completed installation which is in operation. In the following video clip (0:35), you will see the dragline dumping the phosphate matrix into a constructed “pit.” There, the monitor blasts the matrix into fine particles. The slurry pours out of the pit, through the iron grate that is designed to keep large chunks from passing, and into a sump where the slurry is drawn into the pump and transported into the slurry pipeline that goes back to the plant. The video clip is 35 seconds in length, but actually represents about 10 minutes of actual work. In other words, you are seeing it in fast motion!

9.5.7: A Closer Look

Let’s look at this construction in more detail. Looking at the next picture, we have a hole that has been dug, and they call it a “pit.” The pit is bisected by the steel grate, which they refer to as the “pit trap.” To the left of the grate, we have the hydraulic monitor, and this is where the dragline deposits the phosphate matrix. On the right side of the grate, we have the slurry intake pipe. In this photo, the intake has been lifted above the water line. Normally, the intake is submerged below the water line.

Figure 9.5.16: "Pit" showing slurry intake, "pit gun," and the "trap."

Source: The Mosaic Company [90]

In this next photo, we can see more clearly the construction of the pit and pit trap.

Figure 9.5.17: Pit trap construction

Source: The Mosaic Company [90]

The next step in the construction of the pit is to install the monitor. Small dozers are used to construct the pit and to do the grading around it. In this next photo, look to the right of the worker. Do you see the specially outfitted tractor with the winch/crane assembly? This is what they use to move and position the monitors, pit traps, pumps, and related assemblies.

Figure 9.5.18: Installation of the "pit gun"

Source: The Mosaic Company [90]

The next picture provides a wide-angle view of the pit and related equipment.

Figure 9.5.19: Complete view of the monitor in operation

Source: The Mosaic Company [90]

9.5.8: Construction of a New Pit

Now, with this understanding of the individual components, let’s watch a video (8:57) showing the construction and set up of a new pit. Note all of the related activities in the background.

Of course, as mining advances, it becomes necessary to move the pit to the next location. This is a process accomplished over a shift. The next video, which is about 6 minutes in length, captures a complete move over a nine-hour period. Fasten your seatbelts!

The design of a slurry transport system is covered in our materials handling course, MNG 404. Here, it is worth noting a few defining characteristics of the system used at the Mosaic mines.

• 2,000 tph capacity
• 1,500 hp or greater pump motors
• up to 62” diameter pump impellers
• pipelines up to 10 mi long
  • 20” diameter pipe
  • pumps every 1 mi
• slurry
  • 35% solids by weight w/ up to 6” – 8“ size fraction

Finally, we can take a look at the placement of overburden into a previously mined cut.

9.5.9: Reclamation

Reclamation is the final part of the sequence of operations, and as you know, it is an ongoing activity. Mosaic does a stellar job of reclaiming the land, and given the growing conditions in Florida, within months after mining has been completed, the area appears undisturbed and back to its original appearance. Actually, that is not always true – often the reclaimed area is suitable for a higher purpose than the overgrown scrub areas that preceded mining. On their website, they discuss the reclamation activities in more detail.

Figure 9.5.20: Reclaimed land in a Mosaic promotional photo

Source: The Mosaic Company [90]

With that, our case study and the module on surface mining comes to an end!

10.1.1: Construction of the Access

We should talk a little bit about the construction of these means of access. The construction of the drifts, adits, slopes, and ramps does not differ significantly from the unit and auxiliary operations for mining the ore, which we discussed previously. However, these access openings are generally, except to remain serviceable for the life of the mine, unlike many of the other workings that are in use only while the ore in that part of the mine is being exploited. Consequently, extraordinary measures can be justified to ensure the stability of these openings over many years. What are these measures? Reinforced concrete liners, steel arches, and/or additional rock and cable bolts to ensure long-term stability. Some rock types deteriorate when exposed to moisture, and you may shotcrete these surfaces to prevent oxidation and deterioration.

Shaft sinking, on the other hand, may involve some operations and equipment that we didn’t address directly in our study of unit and auxiliary operation. Shaft sinking can be accomplished with a conventional cycle, i.e., drill-blast-muck-hoist, or a continuous cycle using a blind-shaft boring machine. Let’s talk about each cycle.

Previously, we talked about the conventional cycle used in mining as consisting of the unit operations: drill-blast-load-haul. The same unit operations are used to sink a shaft, but the name for loading has changed to mucking and the name for hauling has changed to hoisting. As we sink a shaft into the earth, we cannot load the blasted material with a wheeled loader for example and haul the blasted material off to a dump… obviously! Instead, we use a clamshell mucker to grasp the broken rock and drop it into a large hoist bucket. When the bucket is full, it is hoisted to the top of the shaft and dumped into a waiting truck, where it will be hauled to a dump pile. Hence, the name change for loading to mucking and hauling to hoisting in the conventional cycle for sinking a shaft. It should be noted that the word muck refers to any broken, i.e., blasted, rock, and the word mucking is an old mining term for the operation of loading out muck. Here is a video illustrating the mucking and loading operations in a shaft. This example is for a fairly shallow and small shaft. Larger and deeper shafts involve more complex arrangements, consisting of multiple deck stages and equipment. Nonetheless, this video (2:28) illustrates the basic concept.

During the sinking of the shaft, the auxiliary operation of ground control is generally crucial. The first tens or even a hundred or more feet are driven in relatively poor quality material, e.g., soils and weathered materials that will not stand on their own. In other words, they would tend to fall into the shaft. In areas of past glaciation, the overburden may consist of a hundred feet or more of loosely consolidated rubble. In these areas, the only way to sink a shaft is to freeze the overburden, drill and blast through the frozen material, and then immediately place a liner in the shaft to support the shaft walls. Glaciers made it as far south as Illinois, for example, and so, if you want to sink a shaft down to a coal deposit in the Illinois basin, chances are good that you will have to freeze the alluvial till, i.e., the unconsolidated overburden. Freezing is accomplished by drilling and placing refrigerant pipes around the location of the future shaft, setting up a large refrigeration system, and pumping coolant through the pipes until the ground is frozen. Then, the shaft sinking can begin. Here is a good video (4:37) to illustrate a shaft sinking process in which freezing the overburden is part of the project.

The ground support to maintain the integrity of the shaft walls may consist of rock bolts and wire mesh, but will normally require the use of liners. The liners could be timber, which was used in the past, mortared brick liners, which are also a thing of the past, or most common, concrete liners. In most cases, the concrete liners, a foot or more in thickness, are poured in place. Sometimes, precast liner segments are delivered to the site and set into place, and then a grout is pumped behind the liner to fill the space between the rock wall and the liner. The liners will change over the depth of the shaft depending on the need. If there is a problem with ground water infiltration, steps will be taken to seal the shaft in the water-bearing horizons. This is not as simple as it may seem. At depth, the hydrostatic pressure on that water could reach 500 to 1000 psi or more! To withstand these forces, specially constructed steel liners are used. They may be made out of steel that is ½ to 1” thick, and then welded into place with a concrete ground pumped behind the liner to fill the void between it and the shaft wall. Here’s a picture of such a liner in the shop prior to delivery.

Figure 10.1.2: Steel liner

Michael Marksberry, Compass Minerals

And here is a picture on site. You can see the liners waiting to be installed, along with the gantry for lowering them into the shaft. In this case, they are being set in the first 900’ of the shaft.

Figure 10.1.3 Steel liners waiting to be installed

Michael Marksberry

10.1.2: Revisiting "The Quest for Continuous"

In a previous lesson on “the quest for continuous,” we looked at the motivation for replacing conventional cycles and the equipment for doing so. In the case of shaft sinking, the same principles have been applied. A variety of boring machines has been developed, and consist of a rotating and bit-laced head, a built-in materials handling system to clear the cut material and send it on its way to the surface, and the means for advancing, guiding, and controlling the borer. Likely, there are stages immediately above the borer that are outfitted to allow placement of liners. The following video clip (10:35) provides a detailed look into the functionality of a modern shaft boring machine.

In soft to medium harness rock, the boring machine is the preferred choice, as it can be done faster and at a lower cost than with conventional methods. However, in hard materials, a conventional drill and blast cycle is the only viable choice. This is similar to what we find for continuous mining methods in various deposits. There is one other machine of note for smaller diameter shafts – the raising boring machine.

Small diameter shafts, say less than 10’ in diameter, are required in many applications. They are used for emergency escape hoists, ventilation shafts, and travel ways between levels in a mine. The latter are known as raises, and this is where the name, raise borer, originated. In the good old days, raises and small diameter shafts were driven using conventional cycles. Today, a continuous cycle is used in most cases.

Raise boring requires that the bottom of the raise terminates in an existing mined opening. The first step in using a raise borer is to drill a pilot hole down to the exiting opening. Once completed, a large diameter cutting head is threaded onto a drill steel that connects to a power source on the surface or the upper level. The power source, i.e., the raise borer, provides thrust and rotation for the cutting bits. The cuttings fall to the lower level where they must be loaded and hauled. Take a look at this figure – it makes more sense if you can see it!

Figure 10.1.4: The raiseboring process

Source: Atlas Copco Underground Methods 

Here is a picture of the power unit. Raise borers can excavate an inclined open as well as vertical, and they can bore upward, although that is done with less frequency.

Figure 10.1.5: A raise borer

Source: Atlas Copco Underground Methods 

10.1.3: Odds and Ends

There are a few odds and ends – points that need to be made, but which I haven’t discussed until now. I’ll finish out this lesson by covered these.

The shaft collar is the name given to the point where the shaft intersects the surface, and it is also a structural component of the shaft. It is typically reinforced concrete and it will be either anchored into bedrock or tied into a liner that is anchored in the bedrock. The collar serves an important structural function for equipment associated with the shaft, and it is used to limit surface drainage into the shaft.

A building or structure is associated with the shaft. If there is a hoist, there will be a headframe. Hoists are always used to haul ore up the shaft and out of the mine. The hoist may also include a cage. This is a steel structure, typically with open steel mesh construction, giving it the name “cage” used to transport miners and supplies. The trip to the bottom or back up to the top is known as a mantrip. In deep mines, the cage will likely have multiple floors so that a hundred or more miners can be transported at one time. When it takes 15 to 30 minutes to make the descent, you want to minimize the number of trips! Hoists are of two general types: drum or Koepe (friction hoist). You will learn how to design both in MNG 404, and we won’t go into the differences here, although they are discussed in the text. If there is no need for a hoist, i.e., to transport ore, but only a need to transport people and supplies, then an elevator would normally be used. The elevator is similar in construction to one that you would find in a high-rise office building of 100 stories.

Here is a cross-section of a shaft showing the compartmentalization or multi-function layout. Half the shaft is dedicated to hoisting ore, and there is a skip at the bottom and a skip at the top of the shaft. As one skip is being unloaded at the top, the other is being loaded at the bottom. For reasons that will become clearer later on, this is a Koepe hoist. The other half of the shaft is dedicated to a cage. This cage is connected to a drum hoist. You might want to think about why the Koepe hoist is used in the one case and the drum hoist in the other.

Figure 10.1.6: A cross-section of a shaft showing the compartmentalization or multifunction layout

Source: Michael Marksberry, Compass Minerals

Shaft development practices have changed over the last few decades, with bored shafts being much more common and the "circular" shaft becoming all but universal for large shafts in new mines. In coal mining, about half of all shafts are now sunk by blind-boring methods. It has become common for main shafts up to 16’ to be drilled in this fashion.  In addition, many coal mining companies now use bleeder shafts to simplify their longwall ventilation systems.  These are sometimes drilled using blind-boring rigs. Raise-bored shafts are done as well, but access must be available underground to make this feasible, and the need for a mucking and transportation system underground makes this alternative less favorable. For large shafts, greater than 20’, in coal mines and for most metal and nonmetal mine shafts, conventional shaft-sinking methods are still commonly used.

We’ve talked little about slopes and ramps because their construction differs little from the conventional mining cycle. However, it should be noted that tunnel-boring machines are being used with increasing frequency to drive large slopes of considerable length. The cost of bringing a tunnel boring machine to a site, and setting it up is very high, and can only be justified for large diameter slopes that are greater than a few thousand feet.

The slopes used in today's coal mines are normally driven with two compartments: an upper compartment to accommodate a belt conveyor and a lower compartment containing track that is normally used for personnel and supply transport. The two compartments are separated by a horizontal concrete divider and are supported by bolts, wire mesh, or steel arches. Slopes in coal mines are often horseshoe-shaped, with widths of 17 to 20’ and heights of 13 to 14‘.

The next lesson will address the "building blocks" or basic elements of underground mines. Once we have gained access to the deposit, we are ready to begin development of the underground workings. We'll be introduced to the elements of those workings in the next lesson, and after that, we will be prepared to examine the development of underground mines for the different mining methods.

Before moving on, use these interactive activities to test your knowledge of important terms in this lesson.

10.2.1: Deposit and Spatial Terms

If we are in a horizontal deposit and we mine an opening into the deposit, that opening will have a top and a bottom. We often use those terms. We might say, “this mine has good top,” and we mean that the rock structure at the top of the opening is competent. Or, we might say, “that seam is underlain by clay, and the bottom is terrible,” and we mean that it is difficult for equipment to move around because of the poor condition of the bottom. In a tabular deposit, we frequently refer to the top as the roof, and the bottom as the floor. Conceptually, this aligns with our everyday experiences – if someone refers to the roof or the floor, we know what they mean. Likewise, if we refer to the top or bottom of our mined opening, we intuitively know what is meant. As I indicated, these four terms, top, bottom, roof, and floor, work well for tabular deposits, and are used in coal, trona, potash, salt, and stone mines, for example.

When we move into metal mines, we will not hear those terms. There are a few reasons for this. The mining methods and culture evolved differently in metal versus coal and some of the nonmetals, and they each evolved their own terminology. It’s similar to the situation with the word we use to describe a soft drink. In some parts of the country we would ask for a bottle of “pop.” In another region, we would request a bottle of “soda.” If you go into an area and use the “wrong” term, people may snicker and smile, but they will know what you mean. The same is true for some of these mining terms, except that in addition to smiling, they will likely think that you are a rookie and perhaps are not really a mining person. If, in fact, you are a mining engineer, then that would not be a good thing!

The second reason for the difference in terms relates to the spatial complexity of many metal mines and metalliferous the deposit. In a steeply dipping deposit, for example, the concept of a roof and floor is less useful than hanging wall, footwall, and back. And, of course, there can be crossover in terms. A mining engineer working in a tabular limestone deposit will refer to the top of the opening as the roof, unless they have a background in the metal industry, and then they will refer to it as the back. Don’t despair! As you look at the figures, study the definitions, and read more about the methods, it will become perfectly clear. On with the definitions!

Initially, students are uncertain about which is the footwall and hanging wall. If you have doubts, remember this: these two terms only have meaning if the deposit is dipping at an angle; and pretend the opening in the deposit is a sliding board, and that you are going to slide down – and when you do, your butt will ride on the footwall. Now you’ve got a silly but effective way to keep it straight. Speaking of the footwall, you will notice that the shaft and the workings are located in the country rock and on the footwall side of the deposit, not the hanging wall. There’s a good reason for that. Why do you think that is the case?

All right, onward with the definitions! Here are two more of major importance.

We’ve talked on various occasions about deposits that dip, and so you already have a familiarity with the term. A synonym that is sometimes used is pitching. The terms steeply pitching or a steeply dipping are synonymous. This is true whether it is a seam, vein, deposit, or orebody.

If you’re a surveyor or a geologist, then you are very familiar with the concept of strike. The dip of a deposit can have a profound impact on the type of mine that we develop, but not so for the strike! The strike provides us a direction on the compass to orient our workings, but little more than that for the mining engineer. Now, for structural and economic geologists… they can get excited by the strike of the vein!

In the process of excavating an opening, i.e., mining, we will often leave behind sections of the orebody, for the sole purpose of providing structural support to the overlying layers of rock. These unmined blocks are called pillars, and they are necessary to prevent a collapse of the mined openings. Certain pillars are given specific names, as follows.

Crown and sill pillars are commonly found in mines for steeply pitching deposits, and barrier and yield pillars are commonly found in tabular deposits that are nearly horizontal.

Finally, within this first group, three more terms:

10.2.2: Directional Terms

The directional terms are quite useful and not too difficult to remember. Let me give you an example to illustrate these terms before I give you the formal definition. Imagine that we are all in a big classroom, and the classroom represents the mined out opening. This classroom has rows of chairs and then near the front of the room there is a table on which I lay my books and materials, and then we have a chalkboard on the front wall. I stand behind the table to lecture. There is a door into the classroom on the rear wall. Got the picture?

Ok, let’s imagine that we are mining in the direction of the chalkboard, i.e., we are advancing in the direction of the chalkboard. We’re going to drill holes into the chalkboard, load the holes with explosives, blast, and load out the broken material. We’ve just made the classroom bigger! The chalkboard where we drilled and blasted is known as the face or working face. It is also known as the breast. The act of mining in this horizontal direction is known as breast stoping.

Now, let’s suppose that I’d like to make the classroom higher rather than longer. So, I am going to drill holes into the ceiling, load the holes with powder, shoot them, and load out the broken material. When I advance in this upward direction, it is said to be overhand stoping. Similarly, if I wanted to enlarge our classroom to a lower level, I would drill down into the floor, blast, and load the broken material. When I advance in a downward direction, it is called underhand stoping.

Finally, I can squeeze one last example out of this classroom setting to help you understand the terms inby and outby. These two terms are very useful to state a relative position. Typically, the relative position is between the working face and the entrance to the mine or some portion of the mine. In our classroom that we are pretending is a mine opening, we have a working face (chalkboard wall) and an entrance (the door in the rear wall). Remember the table near the front of the room. If I am standing behind the table, close to the chalkboard, I am standing inby the table, and you are sitting outby the table. If we had a video camera set up in the fourth row back, the students in rows one through three would be sitting inby the camera, and the students sitting in row five to the back of the room would be sitting outby the camera.

This business of inby and outby may seem a little strange, but these words are extremely useful. There is for example, a regulation that prohibits miners from working inby the last row of roof bolts. Or another that allows certain electrical equipment to be used only if it is outby the last open crosscut. Hopefully, with this example, the following terms will be clearer.

10.2.3: Excavation Terms

These terms describe the types of openings that are created to facilitate mining of the deposit. The adjectives primary, secondary, and tertiary are used to characterize some of the terms. This is to indicate the relative importance of the opening in the same sense as saying interstate highways are primary roadways, two-lane streets in a city are secondary roadways, and the alleys between some streets are tertiary roadways. Some of these terms are specific to certain mining methods, and when we study those methods, they will become clearer. I want you to understand all of them, but for right now, focus on the ones that I have highlighted in red.

If you haven’t done so already, this would be a good time to look at the figures.

This figure from the text illustrates the basic infrastructure for a mine in a steeply pitching deposit. Generally, these mines are spatially complex.

Figure 10.2.1: Basic infrastructure for a mine in a steeply pitching deposit

Source: Figure 1.1 from Introductory Mining Engineering, 2nd Ed., Hartman, H.L, and J.M. Mutmansky

The footwall and hanging walls are not labeled, but I am sure you can identify them after our previous discussion. The answer to the question of why the permanent structures such as the shaft are located in the footwall side of the deposit is: as areas are mined out within the deposit, it is likely that some caving will occur, and the caving can be a consequence or cause of fractures propagating through the hanging wall, and potentially ending at the surface. If there were any structures such as a shaft in the hanging wall of the deposit, they would be destroyed by the normal mining operation.

You’ll note the main levels versus the sublevels, and don’t miss the decline. Also, take note of the exploration-drilling program that is ongoing throughout the life of the mine.

I like the next figure because it shows the footwall drift clearly as well as the sublevels and some of the unit operations within the sublevels. The ore passes down to the haulage level are easy to see in this figure, too.

Figure 10.2.2: Sublevel and unit operations

Source: Atlas Copco Underground Methods

Finally, this third figure allows me to illustrate and elaborate on more of the terms. In thick-bedded deposits, it is often necessary to mine the bed in multiple steps known as lifts or benches. In this example, they are taking the seam in two lifts, and you can see the work underway on the benches. Specifically, you should be able to identify the locations where they are breast stoping, underhand stoping, and overhand stoping. You can also see the pillars that are being left behind to support the overlying strata.

Figure 10.2.3: Taking the seam in two lifts

Source: Atlas Copco Underground Methods

Now that we have a common vocabulary to describe the elements of underground mines, as well as an understanding of these basic elements, we are ready to look at the underground mining methods. In the next two lessons, we will look at the methods. I think it is useful to put the methods into a context of their usage within the industry. That is to say, some methods are rarely applied, and others are frequently applied. Let’s finish up this lesson by taking a closer look at this.

10.2.4: Prevalence of Underground Mining Methods

The actual number of underground mines will fluctuate somewhat year-to-year, as a few close and a few open. In recent years, there has been a more significant shift in the number, as several underground coal mines closed due to market conditions. Even though the actual numbers will vary, the proportions are relatively stable. For the purposes of this illustration, we will use underground mines in the U.S., and I’ll round up the number slightly for ease of comparison. There are approximately 600 underground mines in the U.S. Of those, approximately 400 are coal and approximately 200 are noncoal. The noncoal includes metals, industrial minerals, and stone mines.

Every one of the 400 coal mines utilizes the room and pillar method, neglecting the 2 or 3 underground anthracite mines that employ a hybrid method. All of the other underground mining methods are distributed among the 200 noncoal mines. Let’s look at that group in more detail.

The disproportionate number of room and pillar mines is quite noteworthy. Roughly 75% of the underground noncoal mines employ this method. If you add the coal mines into the mix, roughly 90% of all underground mines are using the room and pillar method. Let me ask you a question. If you were to become an expert in the design of one mining method, which would you choose? Well, certainly your career options would be much better if you chose room and pillar! This is not to say that you don’t need to learn about the other methods! You do!!! However, in our curriculum, we do emphasize this method more than the others, and now you know why!

Let’s look at the commodities mined by the three most prevalent methods: room and pillar, cut and fill, and open stoping; as well as the two rarely used methods (in this country) of shrinkage stoping and block caving.

Please be aware that not all commodities are accounted for in this table. There are additional industrial minerals mined by an underground method, which are not accounted for in this data. Despite the very small discrepancies in the totals, the data illustrates accurately the prevalence of the mining methods by commodity.

Review your knowledge of Module 10 terms by engaging with the interactive activities.

10.3.1: Room and Pillar Method

Room and pillar mining is arguably the most important underground mining method in practice today. The majority of underground production comes from room and pillar mines, and the majority of underground mines, by number, employs the room and pillar method! Think about that!

Let’s start out by looking at this sketch of a section in a room and pillar mine. Immediately, you can see that only part of the deposit is mined. Openings are driven in the direction of mining, as shown, and unmined pillars are left in place to support the overlying strata.

Figure 10.3.1 A section in a room and pillar mine

Source: Arch Coal

Remember from our earlier discussion of ground control: when we mine an opening, the weight of the overlying strata must be supported; otherwise, it will cave. As long as the rock layers over the opening are sufficiently strong (think beam), the weight of the overlying members will be transferred to the points where the beam is supported. Those points are the pillars. And from an engineering perspective, it is essential that you do not make the beam too long, because if you do, the beam will fail in the middle, and you will have a cave-in. Just to make it more interesting, you should know that in addition to choosing an appropriate opening width, which is governed by the allowable span of your beam, you also have to worry about the pillar. In some cases, the pillar is not strong enough to bear the weight being superimposed on it, and it will fail. And, in some cases, the pillar will be sufficiently strong, but the layers comprising the floor will not be, and the pillar will push through the floor. Lots of things to think about! This is one reason why, if you are going to be a mining engineer, you will take a course in rock mechanics and cover ground control design in the underground and surface mining courses.

Anyway, back to our sketch of the room and pillar section. The diagram shown is labeled specifically as a coal mine. In fact, it could just as well be salt, trona, lead, and so on; but with some differences that we will discuss. The active mining areas of coal mines are known as sections. In this diagram, you see one working section. This section consists of the equipment and personnel required to conduct the mining activity. From our earlier discussions of unit and auxiliary operations, you will recognize this as a continuous mining operation; and in the U.S., there are no remaining conventional mining sections in underground coal mines.

The mined-out areas in the sketch are given special names, and these may vary depending on the type of deposit that is being mined. There is one term of special significance: the mined-out areas in the direction of mining are known as rooms. Hence, the name of the mining method, room and pillar. Very clever… Typically, the pillars are laid out in this regular checkerboard pattern in coal mines, and now in most other commodities as well. That was not always the case for the noncoal mines. The size, spacing, and even location of the pillars would vary significantly, as would the dimensions of the openings. In those mines, the method was known as stope and pillar. Although you will still hear the term being used, the distinction has largely disappeared, and the term room and pillar is normally applied across all deposit types employing this general method. As a point of interest here, I would mention that mining engineers now recognize that there are serious shortcomings to the somewhat random placement of pillars, resulting in unnecessary ground failures, e.g., cave-ins. As a result of the art and science developed in underground coal mines, ground control approaches such as the pressure arch approach are more generally applied in all commodities, and this results in a more uniform placement of pillars. You will learn more about this if you take a rock mechanics course.

I want to talk about a few more terms. We’ve defined rooms and pillars. The openings driven between rooms are known as crosscuts. Here, they are shown at an angle of 90 degrees, and that is common; but if you are using continuous haulage, such as the flexible conveyor trains that we covered earlier, then you’ll be driving the crosscuts at a different angle of say 60 degrees. Recall also, that the point at which the material is being freed from the deposit is known as the face. In the sketch, you can see five faces. The continuous miner is mining at one face, and a roof bolter is bolting at another face. Alright, there are just a few more terms, and then we fill in some additional detail for the method itself. Let’s look at this plan view of a room and pillar section.

Figure 10.3.2 Plan view of a room and pillar section

Source: adapted from Stefanko, 1983

The sequence of rooms in the direction of mining is known as an entry in a coal mine. They take on the appearance of well-laid-out streets in a city. Indeed, you can stand in an entry and see for quite a distance. If you are in a noncoal mine, you may refer to entries as well, but more likely you’ll call them drifts or headings. What about the sequence of crosscuts? What special name do we assign to that? We don’t assign a special name, and the primary reason is that they do not form a continuous path in the way that rooms do for entries. The reason for that will become clearer within this lesson.

The collection of rooms and pillars shown in this figure form a panel. In this case we have a five-entry panel. Depending on the mining plan this panel could be 10,000’ or more in length, but its width will be determined by the width of the pillars and entries. Three-entry panels are common, as are four and five. There are additional details of note on the plan view.

• There are two different symbols, which appear in the cross cuts: the double line is a solid barrier known as a stopping, and often constructed of concrete blocks; and the single line with a “C” is a flexible curtain, known as a check curtain. Both of these are used to route and separate ventilating air. Recall that you need to provide fresh air to the working places, and then you need to exhaust the air that has become fouled with dusts and gasses.
• The directional arrows are illustrating the direction of the ventilation streams, sometimes called air courses. The air moving toward the face is known as the intake, and the stale air that has already passed the working face is known as the return.
• The thick solid line in entry number three is the conveyor belt.
• The numbered blocks are representing a cut sequence, i.e., block 1 is mined and then block 2, and so on. The cutting sequence will vary by mine, so you don’t need to know this cutting sequence. Simply, it is illustrating that there is a pattern to be followed. This pattern will be designed to optimize productivity and safety.
• The dimensions that are shown will vary somewhat by mine and certainly by commodity; and again, you don’t need to remember these specific numbers.

Finally, in terms of this overview, let’s look at this next figure. Notice that we are no longer representing the entries and cross cuts the same way. We’ve replaced them with a single line. That makes it easier and faster to draw these diagrams. As you look at this figure, you will see some of your newly acquired concepts, including Panels, Sections, Intakes, Returns, and Stoppings. There are also three new terms: overcasts, mains, and submains. Overcasts are yet another type of ventilation control joining curtains and stoppings as controls to route ventilating air. Specifically, overcasts are used to route on type of air over tip of another. It’s similar to a pedestrian overpass to allow people to walk over top of a busy highway. The overcast allows us to route, for example, intake air over top of a return aircourse without mixing the two airstreams. To satisfy your own curiosity, go ahead and trace the airflows in the part of the mine represented in the figure.

Now, on to the two other terms that I really wanted to highlight in this figure: the mains and submains. These are common terms in every coal mine and in some industrial mineral mines, e.g., trona. These words are simply designating their importance in the overall mine plan. The mains serve as the primary means of distributing utilities throughout the mine as well as being the location for the primary transportation and materials handling routes. The submains branch off of the mains to provide these same services to a group of panels, and the panels of course are the location for the active production sections. These word, mains and submains, and sometimes panels, are used as adjectives as well as nouns. The major aircourses supply air for the mine are known as the main intakes and main returns, for example. Main haulage of the mine may be a 72” belt, whereas the belts in the submains may be 60”, for example. It is no coincidence that the mains have more entries than the submains, which usually have more entries than the panels. Mains with seven to eleven entries are common. Often, three or four parallel entries are required to serve as intakes in larger mines, with two or three parallel returns, and another two isolated entries for material handling -- one being a belt entry and the other a track (rail) entry.

Figure 10.3.3 Mine section with two operating units, seven-entry main, five-entry submains, and four-entry panels

Source: Adapted from Stefanko, 1983

We now have a basic understanding of the layout for room and pillar mines, and we know the key terms that are used to describe them. We also know either continuous or conventional production cycles can be employed. With this as a solid foundation, let’s complete the picture with the conditions necessary to use a room and pillar method.

This is an underground method for which you can find mines as shallow as 60’ and at depths of greater than 2500’, and so we can conclude that depth is not a particular defining characteristic for the use of this method. The method does require tabular deposits, as opposed to the porphyry or vein deposits; and further, the method requires that the ore be fairly uniform in quality and thickness. Those are defining characteristics. Deposits with little dip (< 15) are necessary, and less dip, the better. There are rare examples of room and pillar being applied to steeply pitching (dipping) coal seams, but they need not concern us at this time. The rock strength needs to be moderate to strong. The rock needs to be strong enough to allow a reasonable span of opening between pillars. The ore strength on the other hand is not quite as important in the choice of the method. As the ore strength declines, it will be necessary to leave larger pillars, and at some point, that becomes uneconomical. In real-world situations, however, the ore strength is rarely an important characteristic for the selection of this method.

Before reading on, please pause for a minute and think about the characteristics that will lead you to select or exclude the room and pillar method from consideration. Specifically, think about the relationship between that characteristic and the design or operation of a room and pillar mine.

I identified the shape of the deposit as important in the selection of this method, and I said that uniform thickness is desirable. In fact, the thickness can vary by 10 or 20%, and not eliminate room and pillar as a viable method. In some cases, the quality of the ore declines rapidly as you approach the interface between the ore and host rock. In those cases, it is not unusual to leave anywhere from several inches to several feet unmined. In still other instances, the competency of the rock in the immediate roof may be very poor, and in those cases, that material will be mined along with the ore. Yes, that will dilute the run-of-mine product, but the additional cost of doing so, may be less than the cost of the ground-control problem that would result from attempting to leave the “bad” roof layer in place.

I did not say anything about the thickness of the orebody. Room and pillar is used successfully in deposits as thin as 24” and as thick as 100’ or more. It is clear that the orebody thickness is not a defining characteristic of the method itself. However, the mining plan and cycle will be affected as the thickness increases. Consider this: you have a continuous miner with a reach of say 15’. Your seam is 25’ thick. How you are going to mine that seam? Are you going to take 15’ out of the 25’, perhaps down the middle, and leave the remainder? Although there might be an instance in which you would do that, generally you would not invest the capital to access the orebody, and then voluntarily leave a lot of it behind! So, back to the question… what are you going to do?

Why not take it out in layers? That is what we do, and we refer to it as benching. We can do this, and it is frequently done with the continuous or conventional cycles. In some instances, both are used, i.e., the top lift or bench is taken with a continuous cycle and the bottom bench is taken with a conventional cycle. In these thicker seams, three of more lifts may be taken. Take a look at this figure.

It is apparent that the first bench is taken at the top of the deposit. This is the norm for room and pillar mining. For one thing, it is easier to scale and bolt the roof form this first bench. This figure is illustrating a conventional cycle, and we can see a couple of interesting practices. Note the drifting or breast stoping occurring on both the top bench and on the lower bench right side. As a contrast, look at what’s happening on the left side and front of the top bench: underhand stoping. They are also showing some overhand stoping, but, to be honest, I have no idea why! If they were moving upwards in the orebody that would make sense… Just ignore that part of the figure! Anyway, this is a good example of benching used in room and pillar mining.

Figure 10.3.4 Room and Pillar employing a conventional cycle

Source: Atlas Copco

A defining characteristic for the selection of the room and pillar method, as I explained earlier, is a relatively flat lying deposit. What if you meet the characteristics of the unsupported class of methods, except that you have a steeply dipping orebody? You will look more closely at selecting an open stoping or shrinkage stoping method. Next, we will look at these two unsupported methods that are only applicable to steeply dipping deposits that are greater than 50 degrees, and frequently near vertical.  The dip angle of the footwall must be greater than the angle of repose for the broken ore because these methods depend on gravity flow to collection points (draw points).